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ERRATA. 






Page x, line 3, omit 192. 

Page x, line 23, for 303, 304, 306, 307 read 302, 303. 304. 
305. 30 i, 207. 

Page 43, line 39, for November 23 and December 14, 1890, 
read February 11, 1888 

Page 158. line 41, for May 24 read Marcb 24. 

Page 164, line 21, for In figure 79 is given the plan of two 
furnaces forming part of a series of s'x 
read I q figure 79 are given the plans of a 
roasting house containing six long-hearth 
slag-roasting furnaces. 

Paere 210, line 11, for 22 read 7 6. 

Page 214, line 38./or 19*7 pounds, read 18 7 pounds. 

Page 214, line 39, for 34 *S pounds, read 32*7 pounds. 

Page 214, line 39,/or 54*2 pounas. read 51*4 pounds. 

Page 214. line 40. for 72 pounds, read 70 pounds. 

P^ge21 : >. line II, for 72*7 lbs. matte, read 69*9 lbs. matte. 

Page 2lfi, line 30, for Feo, read FeU 

Page 367, line 35, for May 27. 1887, read May 21, 1887. 



THE 



METALLTTKGY OF LEAD 



AND THE 



Desilverizatton of Base Bullion 



BY 

H. O. HOFMAN, E.M., Ph.D. 

ASSOCIATE PROFESSOR OF MINING AND METALLURGY IN THE MASSACHUSETTS INSTITUTE OF 

TECHNOLOGY ; MEMBER OF THE AMERICAN INSTITUTE OF MINING ENGINEERS, 

ETC., ETC. 



/ c~ SECOND EDITION. 




T 

THE SCIENTIFIC PUBLISHING CO. 

27 Pare Place, New York 

1898. 



i 



{ x •; 



V 






Copyrighted, 1892, by 
The Scientific Publishing Company. 



6-/150Q 



TO 

HIS WIFE 

THIS BOOK IS DEDICATED BY 
THE AUTHOR. 



PREFACE. 



In 1870 Dr. Percy published his great work, "The Metallurgy 
of Lead, Including Desilverization and Cupellation," which has 
become the standard book for England and America, and is also 
highly esteemed by the Germans and the French, into whose lan- 
guages it has been translated. It contains an exhaustive discus- 
sion of smelting and desilverizing as carried on in the principal 
European works. During the past twenty years, however, so 
much that is new has been introduced in American practice that 
a book embodying the latest improvements seems almost a neces- 
sity for the student. It is the aim of the present work to supply 
this need. Thus, while European practice is not at all excluded 
from the book, the main object has been to make it a guide for 
Americans, and European methods have been discussed more in 
connection with their applicability in this country than from any 
other point of view. In the subject of the blast-furnace, for in- 
stance, only such features have been brought out as seemed useful 
for America. Reverberatory-furnace practice, on the other hand, 
which has not made much progress as yet in this country, is given 
more in the European way, while the ore-hearth treatment follows 
both English and American methods, which supplement one an- 
other. Three classes of readers have been kept in mind — the stu- 
dent, for whose use the ground principles and many definitions are 
given a prominent place; the metallurgist, who needs minute de- 
tail for his practical operations; and the investigator, who will 
find in the foot-notes the principal references for the subject in its 
various branches. 

To insure the latest data a number of important German works 
were visited in 1890 and the representative American ones in 
1891, and the information obtained has been embodied with the 
name of the authority given, whenever this has been permitted. 
The author desires to thank all who have assisted him in this way 
and otherwise to amplify his notes, gathered through many years 
of practical life. 

II. O. H. 

Massachusetts Institute or Technology, May, 1892. 



PREFACE 

TO THE SECOND EDITION. 



As the first edition of this book only appeared in December, 
it is impossible to do more for the second, which has been called 
for in an unexpectedly short time, than to add a list of errata. 
The author will be grateful to readers who will call his attention to 
any still uncorrected errors, whether of type or otherwise, or who will 
favor him with suggestions through which his work can be im- 
proved in each new edition. A treatise on any technical branch 
of science must necessarily represent to a great extent the per- 
sonal experience and conclusions of the writer. Thus he is very 
liable to emphasize difficulties or advantages familiar to himself 
in a given mode of practice at the expense of others with which he 
has not happened to be brought in contact. It is only by compar- 
ing the experiences and views of others with his own that anyone 
can hope to produce in the end a book of general and lasting 
value. 

H. 0. H. 

Massachusetts Institute of Technology, January, 1893. 



TABLE OF CONTENTS. 



PAGE 

Preface iii 

List of Illustrations ix 



PART I.— INTRODUCTORY. 

Chapter I.— Historical and Statistical Notice. 

§ 1. Introductory Remarks.— § 2. Lead in the United States. 

— § 3. The World's Production 11-13 

Chapter II.— Properties of Lead and of Some of its Compounds. 
§ 4. Lead.— § 5. Lead Oxide.— § 6. Lead Silicates.— § 7. Lead 
Sulphide.— § 8. Lead Sulphate. — § 9. Roasting Lead Sulphide ; 
Reactions between Lead Sulphide, Lead Sulphate, and Lead 
Oxide. — §10. Lead Carbonate. — § 11. Lead of Commerce 14-24 

Chapter III.— Lead Ores. 

§ 12. Introductory Remarks. — § 13. Galenite.— g 14. Anglesite, 
Cerussite. — § 15. Other Lead Minerals 25-29 

Chapter IV.— Distribution of Lead Ores. 

§ 16. Lead Ores of the United States : I. Lead Ores of the Atlan- 
tic Coast. § 17. Lead Ores of the Atlantic Coast. — II. Lead Ores 
of the Mississippi Valley. § 18. The Upper Mississippi Valley. — 
§ 19. The Mines of Missouri. — §20. Other Occurrences. — III. Sil- 
ver-Lead Ores of the Rocky Mountains. § 21. Colorado.— £ 22. 
South Dakota.— § 23. Montana.— § 24. New Mexico.— IV. Silver- 
Lead Ores of the Pacific. §25. Nevada.— §20. Utah.— § 27. Idaho. 
— § 28. Arizona and California 30-1 1 

Chapter V.— Receiving, Sampling, Assaying and Purchasing 
of Ores, FLUXES and Fuels. 
§ 29. Receiving and Weighing of Ores. — § 80. Tl^e Moisture Sam- 
ple. — § 31. The Assay-Sample. — § 32. Genera,] Arrangement of 

Sampling Department.— £ 83. Receiving and Sampling of Fluxes 



vi TABLE OF CONTENTS. 

PAGE 

and Fuels.— § 34. The Assaying of Lead-Silver Ores. — § 35. Pur- 
chasing of Lead-Silver Ores. — § 36. Purchasing of Non-Argen- 
tiferous Lead Ores. — § 37. Purchasing of Fluxes and Fuels 45-76 

PART II.— THE METALLURGICAL TREATMENT OF LEAD ORES. 
§ 38. Classification of Methods 78 

Chapter VI— Smelting in the Reverberatory Furnace. 

§39. Introductory Remarks. — §40. Influence of Foreign Matter. 
— £41. Classification of Reverberatory Methods. — §42. TheCarin- 
thian Method.— § 43. The English Method.— §44. The Silesian 
Method. — § 45. Comparison of Reverberatory Methods 79-109 

Chapter VII.— Smelting in the Ore-Hearth. 

§ 46. Introductory Remarks. — § 47. Influence of Foreign Matter. 
— § 48. Description of Ore-Hearths. — § 49. Mode of Working in 
the Ore-Hearth.— § 50. Treatment of Slags. — § 51. Recovery of 
Flue-Dust by the Lewis and Bartlett Bag Process 110-131 

Chapter VIII.— Smelting in the Blast-Furnace. 

§52. Introductory Remarks. — §53. Lead Slags. — §54. Action of , 
Fluxes and Influence of Foreign Matter. — § 55. Fuels Used in 
the Blast-Furnace.— §56. The Roasting of Ores.— § 57. The Selec- 
tion of a Furnace Site. — § 58. General Arrangement of a Smelt- 
ing Plant. — § 59. The Blast-Furnace and its Accessory Apparatus. 
— § 60. Chemistry of the Blast-Furnace. — § 61. Calculation of 
Charge. 

General Smelting Operations. 

§ 62. Blowing-In.— § 63. Regular Work on the Charging Floor. — 
§64. Regular Work on the Furnace Floor. — §65. Work on the 
Dump. — §66. Irregularities in the Blast-Furnace. — §67. Blow- 
ing-Out. — §68. Furnace Books. — §69. Furnace Assay Book. 

Furnace Products. 

§70. Furnace Products.— § 71. Base Bullion.— § 72. Sampling 
and Assaying. — §73. Speise. — §74. Analytical Determinatioas. — 
§75. Matte.— £76. Roasting of Matte.— § 77. Smelting of Roast- 
ed Matte.— § 78. Nickel and Cobalt- Matte.— § 79. Analytical 
Determinations. — §80. Slag. — §81. Analytical Determinations. 
— § 82. Wall-Accretions. — § 83. Hearth-Accretions or Sows. — 
§84. Furnace Cleanings and Furnace Refuse. — §85. Flue-Dust 
or Chamber-Dust.— § 86. Treatment of Flue-Dust.— § 87. Ana- 
lytical Determinations.— §88. Losses in Smelting.— §89. Cost of 
Smelting 132-294 



TABLE OF CONTENTS. vii 

PAGE 

PART III. -DESILVERIZATION OF BASE BULLION. 
§90. Introductory. 296 

Chapter IX.— Pattinson's Process. 

§ 91. Introductory Remarks. — § 92. Description of Plant and 
Mode of Conducting the Process. — § 93. Luce and Rozan's Process 
(Steam-Pattinson Process.) 298-309 

Chapter X.— Parkes' Process. 

§ 94. Introductory Remarks. — § 95. Outline of Plant and Process. 
— § 96. Receiving Base Bullion. — § 97. Softening Base Bullion. — 
§98. Furnaces.— §99. Mode of Conducting the Process.— §100. 
Desilverizing Softened Bullion.— § 101. Desilverizing Kettles. — 
§ 102. Liquating Apparatus. — § 103. Mode of Conducting the 
Desilverization. — §104. Mode of Liquating. — §105. Refining De- 
silverized Lead. — § 106. Refining in the Reverberatory Furnace. — 
§ 107. Refining in the Kettle.— § 10,8. Moulding from the Refin- 
ing Furnace. — §109. Moulding from the Merchant Kettle. — 
§110. Labor, Fuel, Output of Lead.— §111. Treatment of Zinc- 
Crusts.— § 112. Flach's Process.— § 113. Distillation of Zinc- 
Crusts.— §114. Furnaces.-§115. Condensers.— §116. The Meth- 
od of Working.— §117. Tools.— §118. Results.— §119. Roesing's 
Suggested Improvements. — §120. Comparison of the Two Pro- 
cesses. — §121. Treatment of By-Products. — §122. Softening Fur- 
nace Dross and Skimmings. — § 123. Tin Skimmings. — § 124. Ket- 
tle Dross. — § 125. Refining Skimmings and Polings. — § 126. Rich 
Lead and Metallic Zinc— § 127. Retort Dross and Blue Powder. — 
§ 128. Litharge.— § 129. Old Retorts, Cupel Bottoms, etc.— § 130. 
Table of Desilverization.— §131. General Remarks. — §132. Rela- 
tive Advantages of Parkes' and Pattinson's Processes 310-368 

Chapter XI. — Cupellation. 

§133. Introductory Remarks 369 

A. German Cupellation. 

§134. Characteristics.— § 135. The Furnace.— § 136. Plattner's Cu- 
pelling Furnace. — § 137. Mode of Conducting the Process. 

B. English Cupellation. 

§138. Characteristics.— §139. The Furnace.— §140. Test-Rings.— 
§141. Test-Ring Supports.— §142. The Blast.— §143. TheTools.— 
§ 144. Mode of Conducting the Process. — § 145. Comparison of 
Methods 369-402 



LIST OF JLLTISTKATIONS. 



FIGURES PAGE 

1-6. Quartering 49 

7. Brunton's Quartering-Shovel 50 

8. Pipe Ore-Sampler. 53 

9-10. Brunton's Sampler 54, 55 

11. Sampling-Mill with Brunton's Sampler 56 

12-14. Bridgman's Sampler 57, 59, 61 

15-16. Sampling-Mill with Bridgman's Sampler. 63 

17-18. Bridgman's Laboratory Sampler 65 

19-20. Bridgman's Mixer and Divider 66 

21-24. Reverberatory Furnace at Raibl, Carinthia 85 

25-26. Air-Furnace 90, 91 

27-32. Reverberatory Furnace at Stiperstones, England, 

92, 93, 94, 95, 96, 97 
33-46. Tools Used with the Reverberatory Furnace of Stiperstones . . 98 

47-50. Reverberatory Furnace at Tarnowitz, Silesia 101, 102, 103, 104 

51-53. Scotch Ore-Hearth at Ho wden, England Ill 

54-55. American Water-Back Ore-Hearth 113, 114 

56-59. Moffet Ore-Hearth 115 

60-64. Tools Used with the Moffet Ore-Hearth 116 

65 67. Slag-Eye Furnace of the Lone-Elm Works 121, 123, 124 

68-70. Cooling-Pipes of the Lewis and Bartlett Bag-Process . .125, 126, 12!) 

71. Bag-House of the Lewis and Bartlett Bag-Process 131 

72. Crystalline Forms of Lead-Slags 13? 

73-78. Roasting Furnace with Fuse-Box 163 

79. Plan of Roaster-Building 164 

80-81. Omaha and Grant S. and R. Co., Denver, Col., Plan and Sec- 
tion of Works 1 70, 1 71 

82-83. Montana SmeltiDg Co., Great Falls, Mont., Plan and Section 

of Works 1 7:5. 1 75 

84-88. Omaha and Grant S. and R. Co., Denver, Col., Blast-Fui- 

nace 1 77 

89-95. Globe S. and R. Co., Denver, Col., Blast-Furnace 178, 1 ;••», 180 

86-98. Colorado Iron Co., Denver, Col., Blast-Furnace L82 



x LIST OF ILLUSTRATIONS. 

FIGURES PAGE 

99-120. Montana Smelting Co., Great Falls, Mont., Blast-Furnace, 

183, 186, 192, 197 

121-124. Murray's Tuyere-Box 198 

125-128. Fraser and Chalmers' Ordinary Slag-Pot 199, 200 

129-131. Fraser and Chalmers' Slag-Pot with Roller Bearings 200 

132-133. Terhune's Sectional Slag-Pot 201 

134-139. Murray's Slag-Pot 203 

140-144. Nesmith's Dumping-Car 204 

145. Distribution of Charge on Feed-Floor 228 

146-149. Wall- Accretions in the Blast-Furnace 237 

150-152. Bars of Base Bullion 245 

153. Davies' Converter for Desilverizing Speise 251 

154-157. Pribram Roasting-Stalls 257, 258 

158-163. Montana Smelting Co., Matte-Roasting Reverberatory Fur- 
nace 262 

164. Suspended Sheet-Iron Flue 283 

165-166. Sheet-Iron Flues with Triangular Projections 284 

167-168. Water-Cooled Fiue of Sheet-Lead 285 

169-170. Flue with Freudenberg Plates 287 

171-177. Montana Smelting Co , Flue and Dust-Chamber 288, 289 

178-179. Montana Smelting Co., Blast-Furnace Stack 290 

180-195. Pribram, Luce-Rozan Desilverizing Plant 303, 304, 306, 307 

196-197. Parkes' Desilverizing Plant 315 

198-204. Softening Furnace with Water- Jackets 321 

205-234. Desilverizing Kettles with Brick-Work and Castings 330 

235. Steitz Lead Siphon 343 

236. Lead-Moulding Apparatus for Refining Furnace 343 

237-239. Lead-Mould for Refined Lead 344 

240-243. Lead-Moulding Apparatus for Merchant-Kettle 343 

244-248. Faber du Faur's Retorting Furnace 349, 350, 351 

249-251. Tatham's Retorting Furnace 352 

252-255. Pribram, German Cupelling Furnace 369 

256-260. English Cupelling Furnace 381 

261-269. Test-Rings with Fixed Test 383, 384 

270-272. Fraser and Chalmers' Test-Carriage. 386 

273-274. Lynch's Test-Support 387 

275. (261, 262). Test-Carriage with Swinging Test 383, 388 



PART I. 

INTRODUCTORY. 



CHAPTEE I. 
HISTORICAL 1 AND STATISTICAL 3 NOTICE. 

§ 1 . Introductory Rem arks. — Lead was probably known at a 
very early date. The oldest people of whom we have any record 
— the Egyptians — used it in glossing their pottery, and the abun- 
dance of silver among the ancients suggests the presence of argen- 
tiferous lead ores in many places. We know that lead was mined 
in considerable quantities by the Greeks and Romans. The mines 
of Laurion, in Attica, opened up again by a French company in 
1863, flourished especially in the fifth century b.c. The Romans 
extracted large amounts of ore from the mines in the southeastern 
part of Spain, which had been opened in the third century n.c. by 
Hannibal and which form the main lead district of Spain to-day. 
They also carried on mining operations in England and along the 
Rhine from Bale to Cologne. About the year 1000 a.d. the cele- 
brated German silver-lead mines of the Hartz Mountains, Saxony 
and Silesia, and those of Austria were discovered. 

§ 2. Lead in the United States. — At the present day the lead 
mines of the United States occupy the first place. In this country 
Lead mining dates as far back as the beginning of the seventeenth 

Century, when lead was mined and smelted near Falling ('reck, 

1 Pulsifer, W. H., "Notes for a History of Lead," New York, 1888. 
9 C. Kirchhoff, in "United States Geological Survey: Mineral Re- 
sources of the United States, 1882," and following-, Washington, 1888. 



12 METALLURGY OF LEAD. 

Va. During colonial times lead mines were operated mainly in 
North Carolina, New York, and the New England States, but on 
a small scale and not very successfully. The mines principally 
mentioned are the Washington mine, Davidson County, N. C.;the 
Rossie mine, St. Lawrence County, N. Y.; and the mines near 
Middletown, Conn., and Southampton, Mass. None of them were 
worked continuously, and to-day the East produces hardly any 
lead ore. 

The lead ores of Missouri were discovered by Le Sueur in 1700 
or 1701, and were first worked in 1720; of these the Mine La Motte, 
of Madison County, which is worked to-day, was the first discov- 
ered. Rumors of lead in the upper Mississipi Valley were afloat as 
early as 1766, and in 1788 Dubuque obtained from the Indians a 
grant for a lead mine on the place where the city named after him 
now stands. 

From these two districts the bulk of the lead of the L^nited 
States came until the great mines of the West were opened up in 
1867. The total product of the United States in 1825 was only 
1,500 net tons. This increased steadily until 1848, when 28,000 tons 
were produced. After falling off considerably for a little over 
twenty years, the production of the Mississippi Valley increased 
again, and reached 31,351 tons in 1890. As the total product of 
the United States for the same year was 161,754 tons, this district 
produces 19.36 per cent, of the entire output. All the rest came 
from the Western States and Territories, which produce argentif- 
erous lead ores, while those of the Mississippi Valley are practi- 
cally free from silver. 

Argentiferous lead was first produced, according to O. H. 
llahn, in 1866 or 1867, near Helena, Mont., and at Oreana, Nev. In 
1870 the mines of Eureka, Nev., discovered in 1864, were reopened 
and the treatment of ores was begun. Next came Utah, where 
smelters were erected in 1870, followed by Colorado, which came 
into prominence in 1878. Later Idaho, New Mexico, Arizona, and 
California were added to the list. Colorado is to-day the great- 
est producer. Then follow Idaho and Utah, furnishing togeth- 
er about half as much as Colorado; the three together produce 
about 60 per cent, of all the argentiferous lead of the United 
States. 

S 3. The World's Production.— The world's production in 
1884 l was as follows: 

1 " Mineral Resources of the United States," 1885, p. 264. 



HISTORICAL AND STATISTICAL NOTICE. 13 

Countries. Metric Tons. 

Spain 116,293 

Germany 102,584 

England 40,716 

France 7;500* 

Italy 14,000* 

Greece 8,000* 

Belgium 8,000* 

Austria 11,391 f 

Hungary 1,800* 

Russia 500* 

Sweden 400* 

United States 126,<H)7 

Mexico 1,000* 

Turkey 600* 

Australia..... 5,000* 

South America 2,000* 

Total 446,691 

* Estimated. 

t Including litharge, estimated at 80 per cent. lead. 

The price of common pig-lead in New York averaged for the 
years 1886 to 1890 inclusive per pound avoirdupois 4.32 cents. 
The two other principal lead markets are St. Louis and Chicago, 
where the lead is cheaper than in New York by the cost of trans- 
portation. 



CHAPTEE II. 

PROPERTIES 1 OF LEAD AND OF SOME OF ITS 
COMPOUNDS. 

§ 4. Lead. — Lead has a bluish-gray color; on a freshly-cut sur- 
face it shows a considerable lustre, but loses it quickly when ex- 
posed to atmospheric air. It does not crystallize readily. When 
it is cooled slowly, as in the Pattinson process (§ 91), bundles of 
small, imperfect octahedrons form. Also, when refined lead is 
poured at the correct temperature into a warm mould and is al- 
lowed to cool, fern -like crystalline aggregates appear at the sur- 
face. It is the heaviest of all base metals. Reich's figure, 11.37, 
as specific gravity for pure lead at 0° C. (water at 4° C. being 
unity), is the one generally accepted. The specific gravity of lead 
will vary slightly, according as it is cooled quickly or slowly, ham- 
mered or rolled. Commercial lead has a lower specific gravity 
than 11.37 on account of the impurities contained in it. Lead is 
very soft, especially when allowed to cool and solidify slowly. It 
is harder if cooled quickly and if it contains slight admixtures of 
other metals, such "as copper, arsenic, antimony, zinc, etc. The 
grade of commercial lead is often approximately determined by the 
resistance it offers to scratches with the finger-nail and the facility 
with which it makes a gray streak on paper. Lead is very malle- 
able ; it is rolled into sheets and hammered into foil. In the form 
of filings it becomes a solid mass if subjected to a pressure of 
thirteen tons to the square inch, and liquefies at two and one-half 
times this pressure (Roberts-Austin 2 ). 

Lead is not sufficiently ductile to be drawn into fine wire ; its 
tenacity, according to Karmarsch, is inferior to that of most ductile 

1 Condensed for the most part from Percy, "Metallurgy of Lead," 
London, 1870, pp. 8-93. 

2 Engineering and Mining Journal, November 3, 1888 ; Scientific 
American Supplement, numbers 675, 676, 677. 



PROPERTIES OF LEAD. 15 

metals. It fuses at 325° C. (Le Chatelier 1 ), boils at between 1450 
and 1600° C. -(Camelry and Williams 2 ), but cannot be distilled. 
The latent heat of lead is 5.369 ; the coefficient of cubical dilation 
for 1° C, 0.000089 ; the linear coefficient about one-third of the 
cubical. The specific heat between 10° and 100° C. is 0.0314 ; with 
silver as 100, the conductivity for heat at 12° C. is 8.5, and for elec- 
tricity 10.7. 

Lead undergoes no change in perfectly dry air nor in water 
that is free from air ; its surface becomes, however, dull by oxida- 
tion when it is exposed to the atmospheric air on account of the 
moisture which it contains. Similarly, it is oxidized by water that 
is not free from air. If melted in contact with air it oxidizes and 
becomes covered with an iridescent pellicle, said to be the suboxide, 
Pb 2 ; this gradually changes to the oxide, PbO, and if the heat- 
ing be prolonged sufficiently, the red oxide, Pb 3 4 , is obtained. 
The other two oxides which lead forms are the sesquioxide, 
Pb 2 3 , and the peroxide, Pb0 2 . 

The best solvent of lead is dilute nitric acid. Dilute hydro- 
chloric and sulphuric acids have little or no action ; boiling concen- 
trated hydrochloric acid and sulphuric acid of 66° B. dissolve it 
slowly. Organic acids — acetic, tartaric, citric acids — attack it in 
contact with air. 

§ 5. Lead Oxide, Pb O. — Of the different oxides enumerated, 
the oxide is the one that is metallurgically interesting and of im- 
portance. It is obtained on a large scale as massicot and litharge, 
which have different physical properties. Massicot, an amorphous 
yellow powder, is formed by heating lead on a flat hearth to a low 
red heat, removing the film of suboxide as fast as it forms, and 
oxidizing it to yellow oxide. If the temperature be raised to the 
melting-point — that is, to a bright red heat — and the fused oxide 
cooled, it solidifies as crystalline litharge. On a large scale litharge 
is obtained by cupelling argentiferous lead. It crystallizes in 
orthorhombic octahedrons, and is soft and greasy to the touch. 
While molten it is transparent and orange-colored ; when cold it 
is opaque, and its color varies from yellow to red according to the 
rate at which it has cooled ; quick cooling promotes the yellow, 
slow cooling the red color. Yellow litharge is produced on a large 
scale by allowing it to run from the furnace over an iron plate and 

1 Engineering and Mining Journal, October 11, 1890. 
* Journal of the Chemical Society, xxxv., p. 563. 



16 



METALLURGY OF LEAD. 



chilling it with water, if necessary ; it is thus obtained in small 
lumps. The red, flaky variety is formed by allowing the running 
litharge to collect in front of the furnace in cakes of from one to 
one and one-half tons in weight, and to cool slowly. The inner part 
of a cake will swell up and form flakes of red litharge ; the outer 
and lower parts, having cooled quickly, will remain solid and have 
a yellow color. This swelling is caused by the giving off of oxy- 
gen, which molten litharge absorbs. In solidifying quickly in 
small lumps the oxygen only makes the surface uneven ; in cooling 
slowly in large lumps the outer solid crust obstructs the passage of 
the oxygen. This prevents the inner part from solidifying firmly, 
and causes instead the formation of loose flakes. The flakes and 
lumps are separated by sifting. Both varieties, when ground, have 
a reddish-yellow color. Litharge melts at 954° C. (Honsell x ) ; it 
is a good conductor of electricity when molten. It is volatilized 
at a white heat. It is only slightly soluble in water (1 part in 
12,000 parts), but readily so in nitric acid and acetic acid. 

Litharge is a strong base and quickly corrodes acid furnace 
material, with which it forms a silicate. It is an excellent flux, 
forming fusible compounds with oxides that are infusible alone. 
They do not always enter into chemical combination with it, but 
often are simply held in igneous solution by an excess of litharge. 
Thus fusible mixtures are formed with lime, baryta, magnesia, and 
alumina. The following table shows the proportion of litharge 
required to form fusible compounds with the principal metallic 
oxides: 



1 part of 


Cu 2 


CuO 


ZnO 


Fe 3 4 


Fe 2 3 


MnO 


Sn0 3 


Sb s 3 


3b0 2 


As 2 3 


As 2 6 


Requires 
partB of 


1.5 


1.8 


8 


4 


10 


10 


12-13 


fusible 
in all 
propor- 
tions. 


5 


0.4-0.8 


0.25-1 



Litharge, being easily reduced to the metallic state, forms an 
important oxidizing agent. This is seen by its behavior with S, 
Te, As, Sb, Sn, Bi, Cu, Zn, Fe. They become wholly or partly 
oxidized, and the oxides either volatilized or scorified by the sur- 
plus of litharge, a corresponding amount of lead, which combines 
with any unoxidized part, having been reduced. 

§ 6. Lead Silicates. — Leadpxide and silica begin to combine 
at a temperature where the oxide becomes soft. In fact, it is dis- 



Berg- u. Huttenmannische Zeitung, 1866, p. 106. 



PROPERTIES OF LEAD. 17 

advautageous to raise the temperature quickly if a silicate is to be 
formed. This can be seen in slag-roasting a galena ore to which 
fine sand has been added. If the time at which the roasted ore is 
pasty be shortened and the fusion urged, uncombined silica will be 
found with combined when the roast is decomposed in nitro- 
hydrochloric acid. All silicates that do not contain more silica 
than is required to form the tri-silicate, Pb 2 Si 3 8 (2Pb0.3Si0 2 ), 
are fusible at a low temperature, forming a transparent vitreous 
mass ; the singulo-silicate, Pb 2 Si0 4 (2PbO.Si0 2 ), is as fluid as 
water. If the proportion of silica be raised above that of the tri- 
silicate, the compound becomes less fusible; thus 2Pb09.Si0 2 
gives a porcelain-like mixture, and Pb0.18Si0 2 fritts only to a 
porous mass. All fusible lead silicates are yellow ; they become 
darker in proportion to the quantity of lead they contain. They 
change their color, if they are contaminated with other metallic 
oxides, as can be seen if lead is slagged in a scorifier ; e.g., iron 
colors brown ; copper, green ; manganese, purple-black ; nickel, 
brownish-yellow ; cobalt, blue ; tellurium, yellowish-red, the colors 
growing dark in proportion to the oxide added. 

The lead from silicates is not readily liberated by the ordinary 
reducing agents. Sulphur decomposes the singulo-silicate to some 
extent, but it has less effect on the bi-silicate ; iron sulphide 
throws down some lead, a double silicate of lead and iron being 
the result ; carbon reduces from a bi-silicate part of the lead. In 
order to extract all the lead, it must be first set free from its com- 
bination with silica by a basic flux ; thus metallic iron decomposes 
all fusible lead silicates at a bright-red heat, provided enough is 
added to form a singulo-silicate. 

The singulo-silicate and bi-silicate of lead are readily decom- 
posed by nitric acid, the tri-silicate is not completely decomposed ; 
the more acid the silicate, the less soluble it is. 

§7. Lead Sulphide, PbS. — This occurs native as galenite. It 
is formed by heating lead and sulphur, or lead oxide with an 
excess of sulphur, or by reducing lead sulphate with carbonaceous 
matter. The artificial sulphide has the same properties as the 
mineral. 

The existence of subsulphides of lead (IM>.,S, IM> 4 S) is denied 
by Percy, who shows that lead sulphide and lead can be melted 
together in all proportions, and that the properties of the resulting 
compound will resemble galena <>r lead according to the predomi- 
nance of one or the other compound. Also, if such an. apparent 



18 METALLURGY OF LEAD. 

subsulphide be heated carefully, comparatively pure lead will eli- 
quate and a residue of hard crystalline sulphide remain behind. 

Galena is not as fusible as lead, but it is very fluid when melt- 
ed, and penetrates the fire-brick of the furnaces in which it is 
treated; often a net-work of small veins of bright crystalline galena 
is found in furnace-linings. When melted, galena begins to volati- 
lize without being decomposed, if free oxygen be excluded. On the 
walls of lead blast-furnaces crystals of sublimed galena are of 
common occurrence. Lead sulphide is isomorphous with metallic 
sulphides, as Ag 2 S, Cu 2 S, ZnS, FeS. Such mixtures of sulphides 
are found in lead matte and copper matte obtained in smelting 
sulphide ores. With the electro-negative sulphides of antimony 
and arsenic it forms sulpho-salts. Quite a number of these occur 
as minerals ; x others can be artificially prepared in the dry and wet 
ways. Galena as well as matte is a good conductor of electricity. 2 
Iron decomposes galena better than any other metal. For instance, 
copper that has a greater affinity for sulphur than iron decomposes 
lead sulphide only partially, as it alloys too readily with the liber- 
ated lead, and the cuprous sulphide formed combines with the re- 
maining lead sulphide to matte. Zinc decomposes galena partly, 
but the zinc sulphide formed is so refractory that a separation of 
the liberated lead is not effected; the result is a black, porous mass 
containing particles of lead and galena. The reaction that takes 
place between iron and galena is generally expressed thus: 

PbS-fFe=FeS+Pb, 
and forms the basis of what is called the precipitation or iron- 
reduction process. In reality, however, the iron sulphide retains 
some undecomposed lead sulphide, and Kolte's 3 formula, 

3PbS+3Fe=2Pb-f(PbS+Fe 2 S+FeS), 
might be truer to the actual facts, although he presupposes the exist- 
ence of an iron subsulphide. In decomposing galena in furnace- work 
by means of iron, enough has to be present so as to have lFe for 
lPbS. If less is added, the resulting matte remains too rich in lead; 
if an excess is given, it is wasted. It may in fact be a disadvantage 
in decomposing argentiferous galena, as on account of the affin- 
ity the silver sulphide has for iron sulphide, more silver will go into 
the matte than can be accounted for by the amount of lead pres- 

1 See Dana, " System of Mineralogy," New York, 1880, p. 84. 

8 Kiliani, Berg- u. Huttenmdnnische Zeitung, 1883, pp. 237, 366, 378. 

8 Berg- und Huttenmdnnische Zeitung, 1860, p. 165. 



PROPERTIES OF LEAD. 19 

ent. In addition to having a correct amount of iron present to de- 
compose galena, the temperature is of great importance; the higher 
it is, within reasonable limits, the better will be the decomposition. 
A basic ferrous silicate (4FeO.Si0 2 ) will decompose galena read- 
ily; the singulo-silicate (2FeO.Si0 2 ) shows little effect. In prac- 
tice, the amount of iron that is in excess of that required to flux the 
silica will be available for the decomposition of the lead sulphide. 

Lime as well as baryta has a decomposing action on galena. If 
air has access, the following reaction takes place (Rivot 1 ): 
4PbS+4CaO = 3CaS+CaS0 4 +4Pb. 

If the air be excluded and carbon present, the following occurs 
(Berthier 2 ): 

2PbS+CaO+C=Pb-r-(PbS.CaS)+CO. 

§ 8. Lead Sulphate, PbSO t . — This occurs as anglesite. It is 
formed in roasting lead sulphide (§ 9). Of all metallic sulphates it 
is the only one that is not decomposed upon ignition at a bright- 
red heat; it softens at a white heat and loses some of its sulphu- 
ric acid, forming a basic salt. Silica readily decomposes the sul- 
phate, forming a lead silicate, while the sulphuric acid is driven off, 
being split into sulphurous acid and oxygen. In this way lead sul- 
phate, obtained in roasting a siliceous galena, is decomposed. The 
operation goes by the name of slag-roasting. Lead sulphate is 
a poor conductor of electricity (Kiliani 3 ). Carbon, if present in 
sufficient quantity, reduces the sulphate completely to sulphide at 
a dark-red heat : 

PbS0 4 +2C=PbS+2CO g . 

If there is not enough carbon present, only part of the sulphate 
will be reduced : 

2PbS0 4 -fC=PbS0 4 4-PbS+2C0 2 , and 
4PbS0 4 +2C=3PbSOH-FbS+2C0 8 . 

At a cherry-red heat the resulting sulphate and sulphide will 
react upon each other, as shown in the next paragraph. 

Lead sulphate is only slightly soluble in water and dilute sul- 
phuric acid, more so in nitric acid and solutions of nitrates; it is 
soluble to some extent in sodium hyposulphite, the solubility in- 
creasing with the concentration and the temperature of the solvent 
(Stetefeldt 4 ). 

1 "Traite de mrtallurgie," Paris, 1872, vol. ii., p. 42. 

2 '* Traite des essais par la voie seche," Liege, 1887, vol. ii., p. 580. 

8 Berg- n. Htittenm&nnUche Zeit "/"/, 1888, p. 887. 

4 The LixiviatioD of Silver Ores with Byposulphite Solutions," New- 
York, 1889, p. 81. 



20 METALLURGY OF LEAD. 

§ 9. Roasting of Lead Sulphide; Reactions between 
Lead Sulphide, Lead Sulphate, and Lead Oxide.— The roast- 
ing of galena and making the products react upon still undecom- 
posed sulphide at an elevated temperature is of special interest, as 
one important lead-smelting process, the roasting and reaction 
process, also called the air-reduction process, is based upon it. If 
galena is ground fine and roasted J carefully at a low temperature, 
it will at first be converted into oxide (perhaps only into sub- 
oxide) and sulphur dioxide. Lead sulphide does not oxidize 
readily, and the dioxide will therefore form slowly. As we re- 
quire a low temperature, only part of the dioxide combines with 
the oxygen of the air and forms the trioxide by contact, and 
this combines with the lead oxide, forming sulphate. If lead 
suboxide was present, the sulphur trioxide converts this first to 
oxide. Rammelsberg 2 suggests that some lead sulphide is 
directly oxidized to lead sulphate without passing through the 
stage of oxide. An experiment with pure galena gave to Plattner 
approximately the proportion 5PbO : 2PbSO t ; on roasting a galenite 
from Bleiberg (Carinthia) that contained a small amount of blende 
and pyrite this changed to PbO : PbS0 4 ; with 50 per cent, pyrite 
added the relation was 2PbO : 3PbS0 4 . 

This shows that the relation of lead oxide and sulphate in 
roasted galena depends on the presence of other sulphides. It is 
generally accepted that slow roasting at a low temperature pro- 
duces more sulphate than if the operation be carried on quickly at 
a higher temperature, but, according to Rammelsburg, 3 this is not 
definitely settled. 

If lead sulphide is heated to a strong red heat with lead oxide 
or sulphate, the following reactions will take place : 

(1) PbS+2PbO=Pb 3 +S0 2 . 

(2) PbS+3PbO=Pb 3 +PbO+S0 2 . 

(3) 2PbS+2PbO=Pb 3 + PbS+S0 2 . 

(4) PbS+PbS0 4 =Pb 2 +2S0 2 . 

(5) PbS-f 2PbS0 4 =Pb+2PbO+ 3S0 2 . 

(6) PbS-f3PbSO,=4PbO-h4S0 2 . 

These equations show that with correct proportions (1 and 4) 
all the lead is reduced by the sulphur, which combines with the 

1 Plattner, C. F., "Die metallurgischen Rostprocesse, theoretisch 
betrachtet," Freiberg, 1856, p. 145. 

2 Percy-Rammelsberg, "Die Metallurgie des Bleies," Brunswick, 1872, 
p. 39. 

3 Op. cit.y p. 40. 



PROPERTIES OF LEAD. 



21 



oxygen to dioxide ; if we have an excess of lead oxide (2) or of 
sulphide (3), it remains unaltered. We find something similar the 
case with equations (5) and (6). With too much lead sulphate we 
retain some or all the lead as oxide ; with a surplus of sulphide it 
would remain unchanged. 

§ 10. Lead Carbonate, Pb C0 3 . — This occurs as cerussite. It 
is a poor conductor of electricity. 1 The white lead of commerce is 
a basic carbonate. Lead carbonate is readily decomposed at a 
very low temperature into oxide and carbonic acid. 

§11. The Lead of Commerce : its Impurities and their Ef- 
fect. — In the market we find three kinds of lead : undesilverized 
lead, desilverized lead, and antimonial lead. The first comes from 
the non-argentiferous ores of the Mississippi Valley, the second from 
the refining works which desilverize argentiferous lead (Part III.). 
The third is a by-product of the second. The two soft leads are 
manufactured into sheet-lead and lead pipe ; they are used for 
making alloys, for corroding, and for other chemical purposes that 
require a good grade. The hard lead is used in making type-metal, 
bearings, etc. 

In the subjoined table are given the analyses of the principal 
American brands of lead ; some well-known European makes have 
been added for the sake of comparison. 





Hartz Mountains. 


Pribram. 


Pennsyl- 
vania 
Lead Co. 


Consol. Kansas City 
Smelting and Re- 
fining Co. 


o 


Reduced 

from 
Litharge 


Pattin- 

son 
Process. 


Parkes 
Process. 


Luce- 
Rozan 
Process. 


Parkes 
Process. 


Parkes Process. 


Cu, 
Ag, 
Bi, 
1 Cd, 
Sn, 
As, 
Sb, 
Ni, 
Co, 
Fe, 
Zn, 
Mn, 

s, 
Pb, 
Refer- 
ence. 


0.0600 
0.0028 
trace. 

0.1840 

0.0050 

o'.ooao 

0.0040 

99.7912 
a. 


0.0150 
0.0022 
0.0006 

o'.oioo 

0.0010 

0.0040 
0.0010 

99.9662 
a. 


0.00080 
0.00035 

0.00088 

0.00796 

0.001 00 
0.00092 

99.98480 
b. 


0.0024 
0.0018 
0.0023 

0'.6056 

0.0011 

0.0017 
0.0010 

99.9841 
c. 


0.00007 
0.00042 

0.00051 

trace. 
0.00038 

0.00018 
99.99844 

d. 


0.00022 

0.00020 

0.00308 

trace. 

trace. 

0.00127 

trace. 

0.00178 
0.00075 
0.00021 

99 .99249 

e. 


trace. 
0.0004 

trace. 
0.0004 

o'.oooe 

0.0013 

99.9963 
e. 



a. Zeitschrift fur Berg-. Hiit ten- unci Salinen-Weten in Preuasen, xviii., p. 80S, b. Pri- 
vate Notes. 1890. c. Oeaterreichiache Zeitschrift fur Berg- unci Witten-Weat-n, 1800, D, 497. 
d. Trans. A. I. M. E., iii., p. 322. e. Eng. and Mining Journal, July 14, 1882. /. S«c 8 40. 



1 Kiliani, Berg- u. HiXttenmdnnische Zeitung, 1883, p. 287. 



22 METALLURGY OF LEAD. 

The analyses show that the soft lead in the market is of great 
purity ; the total of the impurities does not exceed 0.02 per cent. 
These are copper, silver, bismuth, cadmium, tin, antimony (ar- 
senic), nickel (cobalt), iron, and zinc (manganese). 

1. Copper. — Copper and lead do not readily form homogene- 
ous alloys. In order to obtain these, the two metals have to be 
melted together beyond the fusing-point of copper and then chilled 
quickly. If such an alloy be heated gradually to the melting- 
point of lead, it is possible to separate part of the lead from the 
copper. This will retain as minimum 0.08 per cent, copper. 1 The 
copper retaining some lead will remain behind as a porous mass. 
In order to remove the copper from the lead an addition of zinc is 
necessary, which will extract it to the practical limit (Parkes Pro- 
cess, § 94). The percentage of copper in commercial lead does 
not interfere with the rolling and other mechanical processes ; it 
appears to protect the lead against the action of sulphuric acid. 
Antimony has a similar effect. 2 If used for corroding or for 
making flint-glass the percentage of copper ought not to exceed 
0.0014 per cent. (Hampe 3 ). 

2. Silver. — What is said about copper in regard to mechanical 
treatment holds good for silver. Small quantities of silver protect 
lead against sulphuric acid. Baker 4 says that 1.70 ounces of silver 
per ton give white lead a reddish tinge, while this is not the case 
with 0.15 ounces per ton. Landsberg 5 gives as minimum 1.03 
ounces for corroding lead. 

3. Bismuth. — A percentage of 0.118-0.352 bismuth makes lead 
hard, 6 somewhat crystalline, and more readily fusible (Hanij)e). 
According to Napier 7 and Bauer, 8 0.10 per cent, bismuth protects 
lead somewhat from sulphuric acid at 20° C, but not at 100° C. On 
white lead, from 0.0045 to 0.0075 per cent, bismuth has no effect 
(Hampe). Endemann 9 states that bismuth favors the corrosion of 

1 See Softening of Base Bullion, § 97. 

2 See below, 6. 

3 Zeitschrift fur Berg-, Hiltten- u. Salinen-Wesen in Preussen, xviii., 
p. 209. 

4 Dingier, Polytechnisches Journal, clxxiii., p. 119. 
6 Wagner's Jahresberichte, 1875, p. 596. 

6 Plattner, Berg- u. Huttenmdnnische Zeitung, 1889, p. 116. 

7 Chemical News, 1880, p. 314; School of Mines Quarterly, vii., p. 97. 

8 Berichte der deutschen chem. Gesellschaft, 1875, p. 40 ; School of 
Mines Quarterly, vii., p. 117. 

8 American Chemist, 1876, vi., p. 457, and Wagner's Jahresberichte y 
1877, p. 422. 



PROPERTIES OF LEAD. 23 

lead, a small black residue remaining behind containing metallic 
bismuth. The only means of removing bismuth from lead is by the 
Pattinson crystallization process (§ 91). 

4. Cadmium. — This occurs only in traces. It protects lead 
from the action of sulphuric acid. 

5. Tin, — Tin makes lead light gray, hard, and decreases its 
fusibility. It is uncommon in market leads. Lead containing 
it is more affected by sulphuric acid than pure lead (Napier, 
Bauer 1 ). The effect in corroding has not been studied. It is re- 
moved by heating the lead to a bright-red heat with access of air ; 
part of the tin collecting on the surface as oxide is first drawn off 
as a powder, and the rest as a slag consisting of stannic oxide and 
lead oxide. (Softening Base Bullion, § 97.) 

6. Antimony (Arsenic). — Even small quantities of antimony 
give lead a grayish- white color, and make it harder and less malle- 
able than ordinary lead. A bar of lead containing some antimony 
will show, especially in the centre, an uneven, moss-like surface. 
Hampe 2 finds that 0.005 per cent, antimony does not harden lead ; 
Heeren 3 states that 0.25 per cent, makes lead hard, but that it is 
still malleable. Lead with 0.1 per cent, antimony is not so easily 
attacked by cold sulphuric acid as pure lead, but more easily by hot 
acid. For corroding, lead may not contain over 0.005 per cent, 
antimony (Hampe, 2 Landsberg 2 ). Antimony and arsenic are re- 
moved, if the lead is heated to a bright-red heat, with access of 
air, as antimoniate and arseniate of lead [Pb 3 .2Sb(As)0 1 ], slagged 
by an excess of litharge. In fact, if tin, arsenic, and antimony are 
present they will be oxidized in the order named and can be 
worked up separately. (Softening of Base Bullion, § 97.) 

7. Nickel (Cobalt). — These hardly occur in market lead. 
Berthier 4 produced a malleable alloy containing from 0.4 to 0.5 
per cent, nickel. Mrazek 5 says that from 1 to 2 per cent, antimony 
favors the entrance of nickel and cobalt into the lead, but they rise 
to the surface when the furnace lead is melted down slow!)', and 
can then be easily skimmed off. 

8. Iron,. — Alloys of lead and iron form under special condi- 
tions, but market lead contains in niaximo 0.07 per cent, iron 

1 See above, p. 22, notes 7 and 8. 
a Loc. cit. 

3 Percy-Rammelsberg, " Die Metallurgie des Bleies,'' p. 49. 

4 "Traite des essais par la voie s<Vhe," Liege, 1847, ii., p. 595. 
8 Berg- u. Huttenmuunische Ze£tung t 1864, p. 815. 



24 



METALLURGY OF LEAD. 



(Reich !), which does not have any effect on the softness and malle- 
ability. Corroding lead ought not to contain over 0.003 per cent, 
iron (Landsberg 2 ). 

9. Zinc. — Zinc and lead can be melted together in varying 
proportions, but they separate again in part on cooling. The 
amount of zinc that lead will retain depends, according to Rossler 
and Edelmann, 3 on the temperature of the lead. The subjoined 
table illustrates this. 



Degrees Centigrade. 


400 


500 


600 


700 


Per cent, of zinc retained 


0.6-0.8 


0.9-1.3 


1.5-2.3 


3.0 



Zinc gives lead a silvery color and makes it so hard that it can- 
not be rolled ; cold and hot sulphuric acid attack it readily. Corrod- 
ing lead should not contain over 0.003 per cent. (Landsberg 2 ). 
Zinc is removed from lead by heating it to a bright-red heat, and 
oxidizing it by admitting air, introducing steam, etc. (See De- 
zincification of Desilverized Lead, § 105.) 

10. Manganese. — This is present only in very small amounts 
and has no practical importance. 



1 Berg- u. Huttenmdnnische Zeitung, 1860, pp. 28, 284. 

2 Loc. cit. 

3 Berg- u. Huttenmdnnische Zeitung, 1890, p. 245 ; Engineering and 
Mining Journal, November 15, 1890. 



CHAPTER III. 
LEAD ORES. 

§ 12. Introductory Remarks. — Quite a number of minerals 
contain lead, but only two or three are found in sufficient quantity 
to constitute lead ore. The lead occurs either as a sulphide or as 
a salt (sulphate, carbonate, etc.). Both are made more or less im- 
pure by other metallic compounds and vein matter. The ores are 
divided into two classes : Sulphide Ores (galenite) and Oxidized 
Ores (anglesite, cerussite, etc.), commonly called Carbonate Ores. 

§ 13. Galenite, PbS; 86.6 Pb, 13.4 S.— This mineral is found 
well crystallized in cubes, sometimes also in isometric octahedrons. 
Crystals are not so often found isolated as in irregular bunches. 
It occurs also in coarsely crystalline to fine granular varieties ; 
crypto-crystalline galena is rare. Galena is found in most of the 
geological formations, but especially in the Silurian, Carbonifer- 
ous and Triassic. It occurs in sedimentary rocks, such as lime- 
stone, dolomite, and sandstone, forming regular beds or impregnat- 
ing the country rock to a greater or less degree, and in fissure and 
metamorphic veins, where it is contaminated with vein matter, 
such as quartz, earthy carbonates, barite, clay-slate, granite, gneiss, 
etc. These are generally removed by a mechanical washing pro- 
cess, and the dressed mineral is then treated by the smelter. The 
mine which contains probably as little galena as any that is worked 
is the one of Mechernich, Rhenish Prussia, where small nodules of 
galena of the size of a pea occur in a soft Triassic sandstone, the 
grains of which, about the size of millet, are slightly cemented 
by a clay or lime cement. The ore contains only 2.5 per cent, of 
galena, and this runs only about six ounces silver per ton. 

The following table shows the rock and formation in which 
some well-known galena deposits occur and the tenor of the ore 
before and after dressing : 



2G 



METALLURGY OF LEAD. 



Locality. 


Nature of Rock. 


Geological Formation. 


? ~ 

it 

a. 


Dressed Ore : 


o 

o 

c 


per 
cent, 
lead. 


oz. sil- 
ver 
per ton. 


Mineral Point, Wis. . 

Rockville, Wis 

Granby, Mo 


Dolomite 

<< 

«« 

Limestone 

a 

Graywack 

Gneiss 

Dolomite 

Graywack slate 
Sandstone 


Lower Silurian . . . 

Sub-Carboni f erous 
Lower Silurian . . . 
(Jarbonif erous. ... 
Triassic 


7 

8.5 

8 

20 
3 
6 
9 
2 


84 

70 
70-77 

71 
37-38 
18-70 
75.5 

64 
56-60 


3.0 
0.3 
1.25 

8 
0.05 
76.5 
17-88 
13.5 
25 
3-4 


1 
1 
1 

2 
3 
4 

5 
6 
7 
8 
9 


St. Joe, Mo 


North of England . . . 
Bleiberg, Carinthia . . 
Pribram, Bohemia. . . 

Freiberg, Saxony 

Tarnowitz, Silesia . . . 
Upper Hartz, Prussia. 
Mechernich, Prussia . 


Lower Silurian . . . 
Archaean 


Triassic 


Sub-Carboniferous 
Triassic 





1. "Geological Purvey of Wisconsin," 1873-1879, iv., p. 382. 

2. Desloge, Trans. A. I. M. E , xviii., p. 262. 

3. Hunt, "British Mining," London, 1884, p. 899; and Phillips "Elements of Metal- 
lurgy," Philadelphia, 1887, p. 566 

4. Oesterreichische Zeitschrift fiir Berg- und Hiittemvesen, 1890, p. 286, 

5. Oesterreichische Zeitschrift fiir Berg- und Huttenioesen, 1888, p. 567; and Oesterreich- 
isches Jahrbuch, xxxix., p. 10. 

6. " Freiberg's Berg- und HUtten-wesen," Freiberg, 1883. 

7. Zeitschrift fiir Berg-, Hiitten- und Salinen-Wesen in Preussen, xxxii., p. 392. 

8. Zeitschrift fiir Berg- , Hiitten- und tialinen-Wesen in Preussen, xxx., p. 131 and 
Private Notes, 1890. 

9. "Bergbau und Huttenbetrieb von Mechernich, 1 ' Koln, 1886. 

Galena often occurs in a very pure state, but it is more gener- 
ally contaminated with other metallic sulphides. These are either 
pyrite, arsenopyrite, chalcopyrite, blende, bournonite, etc., as as- 
sociated minerals ; or silver, copper, zinc, iron, nickel, etc., forming 
isomorphous compounds with the lead sulphide. The associated 
minerals can usually be mechanically separated from the galena, 
but not always. Sometimes the admixture is too intimate, causing 
trouble and loss in the metallurgical treatment. 

Galena is almost always argentiferous. The silver is rarely 
present in the native state ; commonly it appears as isomorphous 
silver sulphide, or as a finely disseminated silver mineral. The 
difference of form is important in connection with concentration *. 
If the silver occurs as isomorphous sulphide, replacing part of the 
lead sulphide, the loss in concentration will correspond to the per- 
centage of lead in the tailings ; if as associated mineral (e.g., 
tetrahedrite), it will be very great, as this mineral, being very 
brittle, is readily crushed to a fine powder, and, being also lighter 
than galena, is carried off on the water. If the dark scum that is 
often seen floating on the water of jiggs, where argentiferous ga- 

1 Raymond, Engineering and Mining Journal, February 11, 1882. 



LEAD ORES. 27 

lena is concentrated, be assayed, the main source of loss in silver 
appears. The tenor of silver in galena ores varies a great deal. 
The galenite from Bleiberg, Carinthia, with 0.05 ounces silver per 
ton, represents probably the lowest amount, and occasional speci- 
mens from Idaho and Schemnitz, Hungary, with 2,042 ounces, the 
highest. The amount of silver contained in galena is often said 
to be dependent on the enclosing rock ; thus crystalline rocks 
would be favorable to a high percentage of silver, and unchanged 
sedimentary strata unfavorable. While many occurrences of ga- 
lena appear to sustain this rule, others again disprove it, so that 
it is not of general application. 

It has often been said, and may sometimes still be heard, that 
coarse-grained galena is poor in silver, while fine granular varieties 
give higher assays, but Malaguti and Durocher x disproved it forty 
years ago. The minerals of tenest associated with galena — such as 
pyrite, blende, etc. — do not generally contain as much silver as the 
galena. 

Percy 2 states that gold is as invariably present in galena as sil- 
ver, but it does not often occur in appreciable quantities. 

§14. Anglesite, PbSO^ PbO, 73.6; S0 3 , 26.4; Pb, 68.3. 
Cerussite, PbCO s ; PbO, 83.5; C0 2 , 16.5; Pb y 77.5.— Carbonate 
ores, using that term in a general sense as embracing all oxidized 
lead ores, occur often in the form of a sand or an earth, bearing 
the name of sand or soft carbonates. In other cases the particles 
of carbonate are cemented together by clay, iron, manganese, or si- 
licious matter, forming compact amorphous or crystalline lumps 
called hard carbonates. 

The minerals are seldom found as originally deposited ; the 
form and composition are more apt to have been caused by chemi- 
cal changes that have taken place since the galena from which 
they resulted was deposited. The sulphate usually formed by the 
oxidation of the sulphido is in most cases the compound from 
which the carbonate has resulted by the action of circulating 
waters holding alkaline or earthy carbonates in solution. 

Anglesite as an ore is rare, as it is not often that galena is 
exposed only to the oxidizing action of air. It often occurs, how- 
ever, with cerussite, and represents the transition between the 
sulphide and the carbonate. This is further illustrated by pieces 

1 Annales des mines, 4th series, 1850, xvii., p. 85. 
1 " Metallurgy of Lead," p. 96. 



28 



METALLURGY OF LEAD. 



of galena found in carbonate ores, which contain in the interior 
anglesite and whose surface is changed to cerussite. Thus these 
two minerals will be always found near the outcrop of galena 
deposits. To what extent the decomposition has progressed de- 
pends on local circumstances. 

The richness in lead of carbonate ores varies a great deal. If 
galena alone has been oxidized, the ores are likely to be rich ; if the 
decomposing action attacked also the country rock, this may so 
contaminate the ore as to reduce the percentage of lead below the 
limit where it pays to treat the ore. The grade of carbonate ore 
is not so often raised by wet concentration as that of galena ores, 
as the losses in lead, and especially in silver, which passes off in the 
slimes, are almost unavoidable. In some instances the carbonate 
ore is first leached with sodium hyposulphite to remove as much 
silver as possible, and then the lead is concentrated in the wet way 
to a rich product. This used to be done at the Old Telegraph 1 
Mine, Ut. Another way is to use Krom's 2 system of dry concen- 
tration, the result being a high grade smelting ore, and tailings 
and dust, to be treated in the wet way. The following table shows 
some very pure carbonate ores from Missouri resulting from cor- 
responding pure galena : 





+3 


Ounces 






Locality. 


0-i 


silver 
per ton. 


Chemist. 


Reference. 


Red Mountain, Col. 


17 


128 


Kedzie 


Trans. A. I. M. E. xvi., p. 581. 


Leadville, Col 


21 


65 


Rolker 


xiv., 287. 


«« < < 


38 


25 


Ricketts. . . . 


xiv., 287. 


Granby, Mo 


65 




Williams. . . 


" " v., 315. 


South West, Mo . . 


72 




Chauvenet . 


Broadhead, " Geol. Survey of 
Mo.," 1874, p. 710. 



The associated minerals undergo a process of oxidation with 
the galena and are generally found again in part in the carbonate 



1 " Tenth Census of the United States," 1880, vol. xiii., p. 415. 

2 Krom : " Commissioner Raymond's Report, " 1876, p. 419 ; Trans. 
A. I. M. E. y xiv., p. 497; Engineering and Mining Journal, 1886, Aug. 
14, Sept. 25, Oct. 23.— Stetefeldt : Eng. and Mining Jour n., Oct. 28, 1876 ; 
May 2, 1885 ; Trans. A. I. M. E., xv., p. 355.— Furman : School of Mines 
Quarterly, hi., p. 127.— Newberry : Ibid., iv. , p. 1. — Heard : Eng. and Min- 
ing Journ., 1886, July 3, Sept. 11.— Dry Recorder: Eng. and Mining 
Journ., 1886, Sept. 4.— Hollister : Trans. A. I. M. E., xvi., p. 1.— Sickel : 
Berg- u. Huttenm. Ztg., 1885, p. 313.— Blomecke : Berg- u. Huttenm. Ztg.„ 
1886, pp. 485, 501, 514. 



LEAD ORES. 29 

ore, although, being more soluble than the lead sulphate, they may- 
be carried away entirely. The silver in oxidized ores is present 
mostly in the form of chloride, although it also occurs in its orig- 
inal form as sulphide and antimonide. There is less likelihood of 
a uniform ratio between lead and silver in a carbonate ore than 
there is in the sulphide ore, as lead sulphate and carbonate behave 
differently with solvents from silver sulphide, chloride, and anti- 
monide. Thus enrichment and impoverishment both in lead and 
silver can be easily accounted for. 

The gold in carbonate ores occurs probably in the native 
state. 

§ 15. Other Lead Minerals. — The following six oxidized 

lead minerals occur often in carbonate deposits, but not in sufficient 
amounts to constitute an ore : 

Pyromorphite, PbCl 2 +3Pb 3 P 2 8 ; 76.4 Pb. — Calcium fluoride 
often replaces in part the lead chloride ; calcium, the lead com- 
bined with phosphoric acid, and arsenic acid, the phosphoric acid. 

Mimetite, PbCl 2 +3Pb 2 As 2 8 ; 69.6 Pb. In the lead arseniate 
the lead is sometimes in part replaced by calcium, and the arsenic 
usually in part by phosphoric acid. 

Vanadi?iite, PbCl 2 +3Pb 2 V 2 8 ; 65.0 Pb. 

Crocoite, PbCr0 4 ; 63.9 Pb. 

Wulfenite, PbM0 4 ; 57.0 Pb. 

Stolzite, PbWoO, ; 45.4 Pb. 

There might still be mentioned about twenty lead-bearing 
sulpharsenites, sulphantimonites, and sulphbismuthites which occur 
in lead deposits, but they are only mineralogical curiosities. 1 

1 Dana, " System of Mineralogy," New York, 1880, pp. 84 to 109. 



CHAPTEE IV. 
DISTRIBUTION OF LEAD ORES. 
§ 16. Lead Ores of the United States.— Lead ores occur in 

many parts of the world. (Table of World's Production, § 3). The 
mines of Spain, Germany, and England furnish the bulk of the 
European product, but it is not the present purpose to deal with 
that branch of the subject. 

The occurrence of lead ores in the United States is best dis- 
cussed under four heads : 
I. The Atlantic coast. 
II. The Mississippi Valley. 

III. The Rocky Mountains. 

IV. The Pacific. 

I. — Lead Ores of the Atlantic Coast. 1 
§ IV. Lead Ores of the Atlantic Coast.— The lead ores of 

the Atlantic coast occur in New York, New England, and Virginia. 
They were worked in former times, and are practically abandoned 
now. 

The Archaean gneiss of New York is traversed by veins of 
galena, which, being free from zinc and iron, have a gangue con- 
sisting mainly of calcspar. The Rossie mines were worked at 
intervals for over fifty years. 

In the New England States, galena, more or less argentiferous 
and associated with other metallic sulphides, occurs irregularly dis- 
tributed in segregated veins in highly metamorphosed palaeozoic 
rocks. 

Virginia 1 has some deposits of galena and blende with carbon- 
ate and silicate of zinc ; the lead, however, is subordinate to the 
zinc. 

1 J. D. Whitney, " Metallic Wealth of the United States," Philadel- 
phia, 1854, p. 382. 

8 Trans. A. I. M. E., viii., p. 340. 



DISTRIBUTION OF LEAD ORES. 31 

II. — Lead Ores of the Mississippi Valley. 

This heading covers two divisions : the lead region of the 
Upper Mississippi Valley and that of Missouri. They have many 
features in common and are often discussed together ; there are, 
however, so many differences in the mode of occurrence of the 
galena, in the associated minerals, and even in the geological 
horizon, that it is advisable to keep them separate. 

§ 18. The Upper Mississippi Valley. 1 — In the southwest- 
ern part of Wisconsin are the important lead deposits of the Upper 
Mississippi Valley, which extend a very small way into the adjoin- 
ing States of Iowa and Illinois. They are principally centred 
around certain districts (diggings), such as Mineral Point and 
Platteville, Wis.; Galena, 111., and Dubuque, la. About ninety 
per cent, of the lead produced comes from Wisconsin. 

The ore is a non-argentiferous galena; it occurs in wholly un- 
disturbed dolomitic limestone of the Trenton period in vertical 
crevices, flat crevices, or as an impregnation. The vertical crevices 
(gash veins) are either thin seams in the rock, a few inches thick 
by several hundred feet long, extending downward for from 
twenty to forty (occasionally one hundred) feet, filled up solid 
with coarse galena; or more commonly they expand in an irregular 
way, inclosing particles of rock, which are cemented together by 
galena. If disintegration has taken place, caves are formed in 
which the galena is found distributed among loose masses of rock, 
calcareous sand, clay, ochre, etc. The horizontal crevices (flat 
openings, flat sheets) are either thin seams of compact ore along 
the bedding planes of the rock, or, if the rock has been disinte- 
grated, have been enlarged, and the galena is found in the same 
form as in the caves of vertical crevices. Combinations of verti- 
cal and horizontal crevices (flats and pitches) increase the size of 
the deposits. Impregnations of certain strata of dolomite occur, 
but they are not frequent. 

The galena (mineral) from the upper beds is pure and rich. 
As depth is gained, the associated minerals — pyrite (sulphur), 
blende (black jack) — increase in quantity and often predominate 
over the galena. Chalcopyrite (copper) is scarce; it is found more 

1 J. D. Whitney, "Metallic Wealth of the United States," p. 403.— 
"Geology of Wisconsin," 1873-1879, vol. ii., pp. 645-752, by M. Strong 
(also in Engineering and Mining Journal, 1878, July 0, fif.); vol. iv., pp. 
877-568, by T. C. Chamberlin.— Trans. A. I. M. E., viii., p. 498, Irving. 



32 



METALLURGY OF LEAD. 



with pyrite and blende than with galena. Secondary minerals are 
not of frequent occurrence. Calcite and barite occur in the lower 
beds. The absence of nickel, cobalt, and arsenic is to be noted. 

§ 19. The Mines of Missouri. 1 — The ore is a coarsely 
crystalline galena, practically free from silver. 

In the Southeastern District, represented by the Saint Joe 
and La Motte mines, galena occurs disseminated through strata 
of dolomitic limestone of the Lower Silurian period, lying almost 
horizontally. The thickness of the lead-bearing " Third Magnesian 
Limestone" varies usually from two to six feet, although it some- 
times goes much higher. The ore as mined at Saint Joe 2 runs 
about 7 per cent, lead and is concentrated to a product of 70 per 
cent. lead. Associated with the galena occurs pyrite containing 
nickel and cobalt. Chalcopyrite, as well as nickel and cobalt 
sulphides with traces of arsenic, is found more at Mine La Motte 
than at the Saint Joe mines. The absence of blende is to be noted. 
The pyrite-bearing galena ores are concentrated (separately from 
the pure galena) to a product called sulphide, which has the fol- 
lowing composition : 



Locality. 


Lead. 


Iron. 


Nickel, 
Cobalt. 


Sulphur. 


Insol. 


Chemist. 


Saint Joe. 1884 

Mine La Motte, 1881. 

it «« a 


21.86 
17.87 
13.34 


16.21 
44.84 


0.61 

4.77 
4.07 


20.37 




3.58 


Setz. 
Neill. 



The dolomitic limestone in which the ores occur contains about 
three per cent, silica; a marked feature is the presence of barite in 
places where the usually disseminated mineral has been concen- 
trated to small sheets. 

The ores of the Central Lead Region, lying between the Osage 
and Missouri Rivers, also occur in dolomite of the Lower Silurian 
period; they resemble those of Wisconsin, but are of no special im- 
portance. 

In the Southwestern Region, which reaches into Kansas, lead 
and zinc ores are worked very extensively at present, especially 



1 Whitney, Op. cit., p. 417.— Broadhead, " Geological Survey of Mis- 
souri," 1874.— Trans. A. J. M. E., iii., p. 116, Gage; v., p. 100, Broad- 
head.— School of Mines Quarterly, ix., p. 74, Kemp. 

2 Trans. A. I. M. E., xvii., p. 659, Munroe ; xviii., 263, Desloge ; 
School of Mines Quarterly, ix., p. 74, Kemp. 



DISTRIBUTION OF LEAD ORES. 33 

around Joplin. The galena occurs in dolomitic limestone of the 
Sub-Carboniferous period, containing layers of chert and bituminous 
matter. When broken, it often emits a bituminous odor. It occurs 
in single or loosely aggregated crystals, also in crystalline masses 
of small dimensions imbedded in the limestone and in the beds of 
chert, the fragments of which are cemented together by a bluish- 
gray, clayey mass. Cadmium bearing blende occurs plentifully 
in two forms, a coarsely crystalline and a granular variety; pyrite 
is subordinate. To be noted is the absence of chalcopyrite and 
barite. Magnesite and calcite are found with the galena and often 
occur in crystalline masses in the dolomite. 

§ 20. Other Occurrences. — The lead deposits of Southwest- 
ern Texas in the Quitman Mountains are as yet of no special im- 
portance. 

III. — Silver-Lead Ores of the Rocky Mountains. 1 

To this division belong the occurrences of argentiferous lead 
ores in Colorado, South Dakota, Montana, and New Mexico. 

§ 21. Colorado. — The deposits of argentiferous lead ores of 
Colorado are more important than any others of the country. 
They are found in veins of Archaean rock and of eruptive rock 
of the San Juan region, and, as sedimentary deposits, in Palaeozoic 
limestone, which is penetrated by igneous rock. The latter de- 
posits are the great metal -producers of the State. 

Boulder County. — On the eastern slope of the Rockies the ores 
occur as veins in Archaean gneiss, which, on account of the indis- 
tinctness of the bedding, Emmons calls granite-gneiss. These are 
not real fissure veins but alterations of the country rock along cer- 
tain planes. The mines are noted principally for the occurrence of 
rich telluride minerals, but in the Caribou district is found rich 
galena with silver minerals, such as stephanite and proustite. The 
Caribou mine contains a massive mixture of rich argentiferous 
galena, chalcopyrite, and blende, occurring in gneiss near a dyke 
of eruptive diabase. 

A shipment of ore made in 1891 contained : 

Pb 44.1* 

Zn 2.r>* 

Fe 9.1* 

Si0 2 8.6* 

Ag 72 . oz. per ton. 

1 S. F. Emmons, "Tenth Census of the United States," 1880, xiii., 
pp. 60-104. 



34 METALLURGY OF LEAD. 

The San Juan Region. 1 — This embraces the southwestern 
part of Colorado, i.e., the counties of San Juan, Hinsdale, Ouray, 
Dolores, La Plata, part of Rio Grande and Conejos. 

Characteristic of the region are immense quartz veins traversing 
older and younger (Triassic) eruptive rocks. The productive ore 
bodies are found in the older massive and brecciated rocks, but 
veins occur also in the underlying gneiss and granite. In the 
neighborhood of Rico (Dolores Co.) and Ouray (Ouray Co.) sedi- 
mentary deposits occur in carboniferous limestones. The deposits 
are mainly argentiferous. The minerals are argentiferous ga- 
lena, silver-bearing gray copper, ruby silver, and native silver. 
Bismuth-silver minerals are frequent ; small amounts of gold are 
found ; blende occurs in considerable quantities. The gangue is 
quartz, kaolinite ; barite is common ; fluorite also occurs. In the 
bedded deposits of Rico iron and manganese are prominent. Near 
the outcrop the ores are often completely changed into a sandy 
carbonate. 

Analyses 2 from two characteristic car-load samples of ore from 
the Red Mountain district by Kedzie are subjoined. 



H 2 (by 
Insol. . . . 


ignition) 


6.75 


11.94 


41.63 


8.65 


Pb 






18.40 




Fe.,0, . . 


17.12 


64.80 


AUO3 .. 






6.08 


13.05 


MnO 


es. 
es. 


Total 


0.59 


1.98 


CaO 


1.70 




ZnO .... 


1.23 




S0 3 


3.87 




co« 


2.39 












99.76 





Ag, ounc 
Au, ounc 


128.00 


68.00 




0.22 


0.10 



Custer County* — Two deposits of this region, the Bassic and 
Bull Domingo mines, differ from other Colorado occurrences. 

1 Emmons, Op. cit., p. 12.— Trans. A. I. M. E., xi., p. 165 ; xv., p. 218 ; 
xvii., p. 261, Comstock ; xvi , p. 804, Emmons ; xvi., p 570, Kedzie ; xviii., 
p. 139, Schwarz. — Proceedings Colorado Scientific Society, March, 1883, 
Hills ; Abstract in Eng. and Min. Journal, 1883, June 9, Emmons. 

2 Kedzie, Trans. A. I. M. E., xvi., p. 581. 

3 Emmons, Op. cit., p. 80.— Grabiel, Trans. A. I. M. E., xi., p. 118. 



DISTRIBUTION OF LEAD ORES. 35 

The ore, which is found in large bodies without any definite 
boundary, forms concentric layers from -J to J inch in thickness 
upon spherical fragments of eruptive country rock. They are 
located near Rosita and Silver Cliff respectively. At the Bassic 
mine the pieces of wall rock vary from § to 24 inches ; the sizes 
most common have diameters ranging from 4 to 12 inches ; the rock 
proper shows only traces of precious metal. The metallic coating 
surrounds the pebbles and bowlders and fills the interestices be- 
tween them. It forms concentric layers. The first one consists 
of sulphides of zinc, antimony, and lead, assays 60 ounces in silver 
and from 1 to 3 ounces in gold, and is dark. The next one is sim- 
ilar, but lighter in color, and assays higher ; the third is blende, 
rich in silver and gold. Chalcopyrite often follows with some 
pyrite. 

At the Bull Domingo mine the occurrence is similar. Around 
a barren nucleus of hornblendic gneiss is deposited a coating of ar- 
gentiferous galenite, followed by siderite, but the ore contains no 
gold. 

The Terrible mine x at Use resembles the other two in some re- 
spects. A porphyry dyke, 127 feet wide, traverses the granite coun- 
try rock for a considerable distance, 87 feet being impregnated with 
crystals of cerussite. As mined the ore carries from 10 to 12 per 
cent, of lead and 1 ounce of silver to the ton. It is concentrated 
and gives an enriched product asssaying 70 per cent, lead and 1.5 
ounces silver. 

Lake County. — Of the sedimentary deposits those of Leadville 2 
are the most important, producing about sixty per cent, of the ar- 
gentiferous lead of Colorado. 

The ore deposits of Leadville occur principally at the contact 
of the dolomitic "Blue Limestone" of the Lower Carboniferous, 
with an intrusive sheet of "White or Leadville porphyry" that 
covers it. They are, however, not confined to the contact, but ex- 
tend into the limestone below. In some cases there is a gradual 
transition from ore to unaltered limestone; in others the ore forms 
pockets and caves in the limestone ; again, the ore has entirely re- 
placed the limestone and is included between two sheets of por- 
phyry. It is exceptional when ore occurs unconnected with the 

1 Private Communication of Messrs. Taylor & Brunton, December 1, 
1890. 

2 S. F. Emmons, "Geology and Mining Industry of Leadville, Col.," 
Monograph XIL, U. 8. Geol. Survey, Washington, 188G, pp 875-588. 



36 METALLVRGY OF LEAD. 

contact surface in the form of irregular masses or running across 
the formation. 

The principal ore is argentiferous galena with its secondary prod- 
ucts of decomposition — anglesite, cerussite, pyromorphite. Silver 
occurs in the altered ore principally as chloride and chloro-bro- 
mide, rarely as chloro-iodide and native, sometimes as sulphuret. 
It has been found that nodules of galena are richer in silver than 
cerussite. Small amounts of gold in the native state have been 
found in limestone, but it is usually associated with pyrite in por- 
phyry. The accessory minerals are blende and its secondary prod- 
ucts, carbonate and silicate of zinc. Arsenic is found as sulphide 
and in combination with iron as arseniate, antimony as sulphide, 
copper as carbonate and silicate, molybdenum and vanadium as 
wulfenite and vanadate of lead and zinc, bismuth as sulphide and 
sulpho-carbonate. Iron and manganese occur as oxides. The 
gangue of the ore consists of silica (as chert, quartz, and combined 
with iron and manganese) and various clays charged with iron and 
manganese. Barite is not uncommon, but irregularly distributed. 
Siderite, pyrite, and gypsum are subordinate. A decomposed 
product, Chinese talc, consisting of a mixture of silicate and sul- 
phate of alumina, occurs at the contact of the white porphyry and 
the limestone. 

The following analysis of carbonate ore by Fluegger l repre- 
sents an average sample of a thousand tons of ore coming from 
every producing mine. It shows the character of the ore mined at 
that time (1881?). 



CO,.... 


... 5.58 


FeO. . . . 


. . 0.89 


MgO .... 


. 3.04 


Au 


. . trace 


PbO.... 


... 25.77 


Fe 2 3 ... 


.. 24.86 


As 


.. 0.01 


Cu 


. . trace 


Ag 


. . . 0.31 


Mn(X . . . 


. . 4.08 


Sb 


. . 0.02 


Zn 


. trace 


Si0 2 .... 


. . . 22.59 


AL0 3 . .. 


. . 3.99 


(KNa)., . . 


. . 0.98 


H s O . .. 


. . 5.58 


S 


... 0.90 


Cab 


. . 2.36 


CI 


. . 0.09 


Total. . 


..101.00 



The composition of the sample in its main features is therefore 
in round figures in per cents.: Pb, 24; Ag, 90.5 ounces; (FeMn)O, 
21.8; A1 2 3 , 3.4; (CaMg)O, 6.6; Si0 2 , 22.6. Guyard, 2 compar- 
ing this analysis with a number of others of that time, 3 says the 
figures of lead, iron, and silica agree with the general composition 
of the smelting charges, but the silver is too high, the common re- 
lation being one ounce of silver to six pounds of lead, while Flueg- 
ger's analysis would correspond to one ounce to every five pounds. 

1 Engineering and Mining Journal, January 8, 1881. 

2 Emmons, Op. cit., p. 620. 

3 Emmons, Op. cit., pp. 621-625. 



DISTRIBUTION OF LEAD ORES. 



37 



The varying proportion of silver and lead in different districts is 
shown by the following figures of Rolker 1 : 



Locality. 


Mine. 


Dry 

weight, 
tons. 


Ounces 

silver 

per ton. 


Per cent. 

of 

lead. 


Relation 

of 
Ag: Pb. 


Fryer Hill 

Carbonate Hill . 
Carbonate Hill . 
Iron H.ll 


Chrysolite 


10,561 
6,315 

4,794 
152,457 


65.45 
39.00 
25.00 
15.00 


21.45 
16.10 
38.45 
18.60 


1 :6.5 
1 :8 
1 :32 
1 :26 


Evening Star 

Morning- Star 

Iron & Silver Mg. Co. 



The table shows that Fryer Hill ores run higher in silver than 
either Carbonate Hill or Iron Hill ores, and that the latter two 
give a lower-grade base bullion. 

These figures refer to the time when mining was carried on en- 
tirely in parts of the ore bodies that had been oxidized by surface 
waters. The ores, just as they came from the mine, could be eco- 
nomically smelted in the blast furnace. Now every plant has a 
number of roasting furnaces, thereby showing how, as the explora- 
tion extends deeper, the unaltered minerals, galena, blende, and py- 
rite are encountered. Unfortunately, at the same time, less silver 
is found. This transition from carbonate to sulphide ore is often 
very rapid. On Iron Hill, according to Blow, 2 the ore changes 
suddenly from a body of fine oxidized smelting ore to a close- 
grained sulphide ore consisting principally of zinc and iron sul- 
phides with sulphide of lead in small quantities. The following 
analyses show the character of some of the sulphuret ores of Lead- 
ville : 





Galena. 


Mixture. 


Galena. 


Mixture. 


Lead Ore. 


Zinc Ore. 


Minnie 
Mine. 


Minnie 
Mine. 


Moyer 
Shaft. 


Moyer 
Shaft. 


Col. Sellers 
Mine. 


Col. Sellers 
Mine. 


Pb 


72.65 

5.66 

1.60 

15.66 

41.50 

trace 

4.12 

' 1 " 


50.86 
12.86 

9.30 
24.50 
11.50 
trace. 

1.88 

"l" 


44.0 
13.0 
11.0 
30.0 
14.0 

2 


15.0 
24.0 
16.0 
40.0 
11.0 

"2" 


27.40 

25.00 

6.00 

35.00 

26.90 

trace 

3.00 

2.00 

3 


10.70 

24.50 

16.60 

40.00 

54.30 

trace 

3.40 

2.10 

3 


Zn 


Fe 


S 


Ag., ounces. 
Au., ounces. 

Insol 

HLO 

References . . 



Rt-ference8 : (1) Trans. A. I. M. E. y xiv. 
xviii., p. 173, Blow. 



p. 189, Freeland ; (2) xiv., p. 288, Rolker; (3) 



1 Trans. A. I. M. E., xiv., p. 287. 
a Trans. A. I. M. E., xviii., p. 170. 



38 



METALLURGY OF LEAD. 



The decrease in silver and lead and the increase in zinc with 
increasing depth have forced the question of concentration upon 
the mining companies. According to Ihlseng 1 several large and 
some small concentrating plants are enriching ores that run from 
6 to 8 ounces in silver and from 6 to 10 per cent, in lead, saving 60 
per cent, of the silver and 70 per cent, of the lead, 7 tons being 
concentrated to 1 ton. 

Taylor and Brunton 2 report that the Colonel Sellers mill, con- 
structed by them, concentrates from 75 to 90 tons of sulphide ores, 
very free from gangue, and containing less than 10 per cent, lead 
and 9 or 10 ounces silver, into heads running 55 to 60 per cent, lead 
and tailings under 2 per cent, lead at a cost of 70 cents per ton. 
The following analyses by Kellar show the character of the prod- 
ucts obtained from coarse-grained ore. 





Crude Ore. 


Blende. 


Galena. 


Pyrite. 


Si0 2 


0.960 
16.185 
22.951 
19.246 

1.664 
trace, 
trace. 
39.240 

7.5 


1.090 

3.525 

10.864 

47.522 

1.477 

37* 040 
3.0 


0.120 
79.958 
2.303 
1.734 
0.154 

15*. 760 
19.8 


0.190 
9.592 

39.431 
3.043 
0.406 

trace. 

trace. 

48.196 
5.2 


Pb 


Fe 


Zn 


Mn 


Cu 


As 


S 


Ag, ounces 





Park County. — The deposits of this county have been devel- 
oped in Archaean and Palaeozoic rocks. The Mesozoic porphyries 
form dykes in Archaean and intrusive sheets in Palaeozoic strata. 
Prominent are the deposits of Mount Lincoln and Mount Bross. 
The ores are argentiferous galena with sulphate and carbonate of 
lead and chloride of silver ; pyrite, more or less decomposed, and 
associated with oxides of manganese, is also found, coloring the 
clay of the gangue ; barite is of common occurrence. The depos- 
its of Mount Lincoln form irregular bodies in the blue limestone, 
and are found generally near its upper surface ; those of Mount 
Bross occur rather in the mass of the limestone near the dykes of 
porphyry. The character of the ore resembles, therefore, that of 
Leadville. 



1 " Report of State School of Mines," Golden, Col., 1887, p. 44. 

2 Engineering and Mining Journal, May 8, 1886. 



DISTRIBUTION OF LEAD ORES. 



3£ 



Pitkin County. — The Aspen 1 deposits of silver-bearing ores 
have lately come into great prominence, very little ore having been 
shipped before the advent of the railroad in 1887. They occur in 
the same geological horizon as the Leadville deposits, viz., the 
Lower Carboniferous limestone, but they are not found in imme- 
diate contact with the eruptive rock. They occur principally 
between two varieties of limestone, the blue (a pure carbonate of 
lime) and the brown (dolomite), which is traversed by numberless 
small veins containing iron salts. The brown limestone breaks 
up by oxidation of the iron into small pieces, and is called " short 
lime." The ore is a limestone impregnated along certain fissures 
with fine-grained argentiferous galena. The main contents of 
silver comes, however, from associated silver minerals, such as 
polybasite and stephanite. Near the outcrop the ore is some- 
what decomposed, consisting of barite, carbonate of copper, oxides 
of iron and manganese, calcite, dice-shaped fragments of dolomite, 
and galena. 

The character of the ore is shown by the following analyses 
furnished by Taylor and Brunton : 2 



BaS0 4 .... 


0.5 


5.5 


9.1 


11.7 


13.0 


15.3 


24.2 


26.0'26.2 


29.5 


30.0 


36.9 


40.0 


SiO a 


6.5 


15.4 


9.2 


20.0 


19.7 


27.9 


17.3 


40.0i44.9 


14.3 


7.0 


22.3 


19.0 


CaO 


7.6 


27.5 


29.8 


19.0 


15.7 


14.2 


13.7 


3.0 1.8 


7.8 


11.1 


12.9 


12.5 


Fe 


9.0 


5.4 


5.5 


4.0 


6.3 


6.5 


1.5 


9.410.6 


2.5 


7.3 


4.8 


6.5 


Pb 


19.3 








1 8 




4 7 


8.0.... 


16 


1 4 






Ag, ounces 


7.5 


18.0 


24.0 


15.0 


26.4 


23.0 


31.0 


25.0 32.0 


69.0 


26.0 


32.037.0 


H 2 


9.75 


3.0 


5.5 


5.8 


7.0 


9.0 


5.0 


9.010.0 


3.0 


9.0 


10.8 3.0 



Summit Count;/. — In the Ten-Mile district the ores occur in 
Upper Carboniferous limestone and in the sandstone above. The 
typical mine is the Robinson. The ore is a rich argentiferous 
galena associated with pyrite and blende. It occurs near the sur- 
face of the limestone, in some cases extends into it, and in others 
penetrates it entirely. 

§ 22. South Dakota. 3 — The occurrence of lead ores is confined 
to two small camps in the Black Hills, Galena and Carbonate. 
Argentiferous galena and carbonate ore occur in shoots in calca- 
reous Potsdam rock, which overlies the upturned Arclnean slates. 
These shoots arc; especially frequent where the porphyry cuts 



1 S. F. Emmons, "Proceedings of the Colorado Scientific Society ; 
Henrich, Traits. A. I. M. K., xvii., p. L66. 

2 Letter of December 1, 1800. 

8 Carpenter, Trans. A. 1. M. E., xvii., p. 582. 



40 METALLURGY OF LEAD. 

through the rocks and where it overlies them. Associated with 
galena, pyrite and blende are found. The ore has to be closely 
sorted, so as to run 20 ounces in silver and 50 per cent, in lead. 

§ 23. Montana. — The argentiferous lead ores of Montana are 
unimportant in comparison with the great copper deposits. They 
occur as metamorphic veins in crystalline rocks or as irregular ore 
bodies in limestones. According to Lindgreen, 1 in the neighbor- 
hood of Helena occur veins from one to ten feet in width, carrying 
in a gangue of quartz, galena, blende, chalcopyrite, pyrite and 
arsenopyrite. The galena assays up to 90 ounces in silver, the 
blende from 10 to 12 ounces in silver, the pyrite from 0.2-0.3 
ounces in gold, the arsenopyrite 1 ounce in gold ; the chalcopyrite 
carries more silver. 

The smelting works at Glendale, Helena, and Great Falls 
obtain a large proportion of their ores from Idaho, especially from 
the Cceur d'Alene District. 

§ 24. New Mexico. — Although New Mexico produces much 
dry-silver ore, lead-silver ores are not of frequent occurrence. The 
principal ones are found in the Magdalena Mountains. 2 The de- 
posits occur in limestone and porphyry and are from 4 to 40 feet 
wide. The principal mines are the Kelly, the Juanita and the 
Graphic. The ore averages 25 per cent, lead and runs very low 
in silver — 8 ounces to the ton. 

IV. — Silver-Lead Ores of the Pacific. 3 

Argentiferous lead ores under this head occur in Nevada, Utah, 
Idaho, Arizona and California. 

§ 25. Nevada. 4 — The production of argentiferous lead in Ne- 
vada has greatly diminished in the last twelve years. Although 
31,063 tons of lead were produced in 1878, in 1890 the production 
had sunk to 1,994 tons. The ore comes principally from two 

1 "U. S. Geol. Survey: Mineral Resources of the U. S. f " 1883-84, 
p. 422. 

2 Silliman, Trans. A. I. M. E., x., p. 425; Prince, Eng. and Mining 
Journal, 1890, November 20. 

3 G. F. Becker, < Tenth Census of the United States," 1880, vol. xiii., 
pp. 5-59. 

4 J. S. Curtis, "Silver Lead Deposits of Eureka, Nev.," Monograph vii., 
"IT. S. Geol. Survey," Washington, 1884.— A. Hague, "Geology of 
Eureka Mining District," monograph, "U. S. Geol. Survey," in prepara- 
tion. 



DISTRIBUTION OF LEAD ORES. 41 

mines, the Richmond mine and the Eureka Consolidated mine, of 
the Eureka District. It occurs in irregular chambers in the Pros- 
pect Mountain Limestone of Ruby Hill, which is a compact mag- 
nesian limestone of the lower horizons of the Cambrian. In the 
upper part of the larger chambers it is in a loose state ; in the 
lower ones it is more compact. It is an argentiferous and aurifer- 
ous carbonate containing both anglesite and cerussite and a very 
small amount of galena ; mimetite and wulfenite are often found. 
A remarkable feature is that at a depth of 1,300 feet the ore is 
still oxidized, and the regular sulphide from which it originated 
has not yet been reached. The associated minerals are pyrite, 
arsenopyrite and blende, with their oxidized products. The gangue 
accompanying the lead ore is principally limonite, which colors it 
in different shades of yellow, red and brown ; lime, magnesia, 
alumina and silica occur only in subordinate quantities. The char- 
acter of the ore is shown by the following analysis made by F. 
Claudet 1 in 1878 : 



PbO . . . 
CuO . . . 
FeO . . . 
ZnO . . . 


... 35.65 
... 0.15 
... 34.39 
... 2.37 


Mn 2 3 .. 
As 2 5 .. 

Sb 

so 3 


.. 0.13 
.. 6.34 
.. 0.25 

.. 4.18 


Si0 3 .... 
AL0 3 .. 
CaO ... . 
MgO .... 


.. 2.95 
. . 0.64 
.. 1.14 
. . 0.41 


HoO+C0 2 10.90 
Ag+Au... 0.10 


Total, 100.52 



27.55 ounces silver per ton ; 1.59 ounces gold per ton. 

The average tenor of lead, silver and gold is lower than the 
analysis shows, viz. : 15 per cent, lead, 23 ounces silver, 0.72 ounces 
gold. 

To be noticed is the large amount of arsenic, which causes the 
formation of the unwelcome speise in smelting. Molybdenum has 
not been determined. Silver is present as chloride and sulphide ; 
gold, probably finely divided, in the native state. 

§ 26. Utah. 2 — The argentiferous lead deposits of Utah form 
more or less regular bodies in limestones or at the contact of lime- 
stone and eruptive rock. The ores are carbonate with anglesite 
strongly prevailing ; the secondary minerals reach down to consid- 
erable depth. Small amounts of galena occur. A very few of the 
mines have been worked down to the sulphuret ores. Others which 
produced only siliceous ores low in lead are running with greater 
depth into bodies richer in lead. 

In Beaver County is the celebrated Hornsilver mine that closed 

1 Curtis, Op. cit., p. 60. 

8 D. B. Huntley, " Tenth Census of the Unit.nl States," 1880, vol. xiii., 
pp. 405-489 ; Hollister, Trans. A. I. M. E., xvi., p. 1. 



42 METALLURGY OF LEAD. 

down in 1885. Prospecting has opened up low-grade sulphide 
ores. The deposit forms a chimney between limestone and rhyolite. 
Anglesite and cerussite are the lead minerals ; galena is found in 
small quantities ; silver occurs as chloride, sulphide and sulpharsen- 
ide ; the gangue is calcite, quartz, barite and ferruginous clay. The 
subjoined analyses 1 by S. B. Newberry, made in 1879, show the 
character of the ore : 

Si0 2 15.17 47.95 

BaS0 4 0.49 2.71 

PbS0 4 74.51 28.80 

Fe 2 3 4.80 

A1 2 3 1.71 12.55 

Sb 2 S 3 0.37 

As 2 S 3 1.12 

(CaMg)O 0.50 

C0 2 0.62 

Ag, ounces 78.33 

The average tenor of lead and silver in 1882 and 1883 2 was 
37.80 and 36.83 per cent, lead and 34.2 and 27.15 ounces silver, 
which is considerably lower than the analysis of 1879. 

The ores of the Tintic district (Juab Co.) are rather milling 
ores (Eureka Hill and Bullion mines), but they are smelted at Salt 
Lake City with lead ores. 

In Salt Lake County are the mines of Bingham and the Little 
and Big Cottonwood Canons. The Old Telegraph 3 mines of L'pper 
Bingham Canon are among the oldest of the territory. The ore 
forms an irregularly mineralized zone between limestone and 
quartzite and sometimes porphyry. The limestone and porphyry 
are more or less decomposed near the deposits. The lead ore is 
carbonate containing more or less silica, ferruginous clay, and 
some galena. The siliceous ore is an oxidized siliceous pyrite, the 
quartz having become ochreous, spongy and brittle ; the pyrite yel- 
low, red or black ochre. 

The following analysis made by Wuth 4 in 1876 shows the char- 
acter of the ore : 



PbC0 3 

PbS 


50.43 

15.02 


Fe 2 3 

CuS 


3.78 

0.67 


FeS 


7.37 



Si0 2 12.47 

A1 2 3 3.01 

CaC0 3 3.64 

MgC0 3 0.26 

CaS0 4 3.04 



H 2 0.19 

Ag, ounces 21 . 14 

Sb trace 

As trace 

Co trace 



1 Huntley, Op. cit., p. 466. 

2 "U. S. Geol. Survey : Min. Resources," 1883-84, p. 417. 

3 Lavagnino, Trans. A. I. M. E., xvi., p. 25. 

4 Huntley, Op. cit., p. 413. 



DISTRIBUTION OF LEAD ORES. 43 

The first-class shipping ore in 1888 contained between 30 and 
50 per cent, lead, 10 and 12 ounces silver and 0.048 ounces gold 
per ton ; the second-class ore is concentrated and runs a little 
higher in lead than the first-class. The ore of the Brooklyn mine, 
lying east of the Old Telegraph mine, is similar in character ; first- 
class ore, 40-50 per cent, lead, 10 ounces silver per ton. Twenty 
tons of low-grade ore are concentrated into 4 to 5 tons of shipping 
ore. The Cottonwood Canons contain the celebrated Emma and 
Flagstaff mines, which stopped producing about twelve years 
ago. 

A number of other mines might be mentioned, but those 
described represent sufficiently the character of Utah silver-lead 
ores. 

While a good deal of Utah ore goes east to be smelted, much 
ore from Idaho comes down to Salt Lake City to be treated there. 
Just as Denver and Pueblo form the main smelting centres east of 
the Rocky Mountains, so Salt Lake City is the ore market of the 
West. 

§ 27. Idaho. — The argentiferous lead ores of Idaho come 
principally from two districts, viz., the Wood River in the south- 
east and the Coeur d'Alene in the north. 

In the Wood River region limestone, slate and granite are the 
chief rocks, rich argentiferous galena with its secondary minerals 
the principal ore. The deposits seem to occur chiefly in lime- 
stones, but little or nothing has been published about them. 
Kirchhoff l says that the ore from the Hailey and Bellevue average 
60 per cent, lead and from 80 to 100 ounces silver. Blake 2 writes 
that the Minnie Moore and the Queen of the Hills furnish concen- 
trates assaying 65 and 62 per cent, lead, 80 and 65 ounces silver 
per ton. 

Some ore is smelted near Ketchum, but large amounts are 
shipped to Montana, Colorado, Utah and California. 

The Coeur cfAlene region. — The rocks are, according to Clay- 
ton, 3 quartzite, magnesian, shale and schist, which have been much 
folded. The deposits form fissure veins varying in width from a 
few inches to twenty feet. The ore, a galena with its products of 

1 "United States (xeol. Survey: Mineral Resources of the U. S.," 
1887, p. 100. 

2 Engineering and Mining Journal, November 29, 1890. 

8 Engineering and Mining Journal, November 23 and December 14, 

1890. 



44 METALLURGY OF LEAD. 

decomposition, is generally concentrated before it is shipped 1 to a 
product assaying from 60 to 65 per cent, lead, and from 25 to 30 
ounces silver per ton. 

§ 28. Arizona and California. — Very little lead is mined in 
Arizona or California. The smelting industry of Arizona is repre- 
sented by the Benson and Tombstone works, but most of the ore 
goes to outside smelters. The lead produced in California comes 
principally from the Selby smelting works at San Francisco, the 
principal one of the Pacific coast. 

1 "United States Geological Survey : Mineral Resources," 1887, 
p. 108 ; Clement, Engineering and Mining Journal, July 25, 1891. 



CHAPTER V. 

RECEIVING, 1 SAMPLING, 1 ASSAYING AND PURCHAS- 
ING OF ORES, FLUXES AND FUELS. 

§ 29. Receiving and Weighing of Ores.— Ores arrive at the 

smelting works either in a loose condition or in sacks. The gross 
weight is taken on platform scales before unloading from the cars. 
The ore is marked, when unloaded, by a shingle 2 by 12 inches, 
with the running lot number chalked on it. The entry is made in 
the receiving-book, a form of which is given below. 

Ores contain a varying percentage of moisture ; those arriving 
direct from the mine are often quite wet or, in winter, frozen solid. 
After being weighed, the ores should be at once unloaded and the 
moisture-sample taken. If the ore is in sacks, these are emj;)tied and 
weighed immediately, before they can become lighter by exposure 
to the air. They are then dried and tied up in bundles. Generally 
speaking, it is best not to receive an ore at all until the operations 
of weighing and unloading it, and taking out and weighing the 
moisture-sample, can be performed in quick succession. If this is 
not possible, the weights will have to be taken after unloading, 
thus causing an extra handling. 

§ 30. The Moisture-Sample. — It is difficult to obtain an ab- 
solutely correct moisture-sample, but a fair average may be arrived 
at which will satisfy the seller as well as the buyer. A correct 
moisture-figure is necessary in making up the charges for the blast 
furnace, as the analytical data from which they are figured refer to 
dry ore, and a correct allowance must be made for the moisture. 

If the ore arrives loose, the surface which has been exposed 
to the air will be drier than the middle part, and this again less 

1 Austin : Engineering and Mining Journal, 1882, July 22, August 5, 
26, Sept. 16. Low : Ibid., Sept. 24, 1881 ; Feb. 11, 1882. Reed : School of 
Mines Quarterly, iii., p. 253; vi., 351. Brunton : Trans. A. I. M. /•.'.. 
xiii., p. 639. Glenn : Ibid., xx. Hodges : Engineering and Mining Jour- 
nal, 1891, Sept. 5. Bridgman : Trans. A. I. M. E. f xx. Johnson : Engi- 
neering and Mining Journal, 1892, Jan. 10, 23. 



46 METALLURGY OF LEAD. 

dry than the real average ; if in sacks, it may show a greater 
irregularity, as more surfaces are exposed to the drying influ- 
ences of the air. With ore of uniform size, arriving loose, the 
sample is obtained by taking a little from the shovel at regular 
intervals while the ore is being unloaded. Or, if it is discharged 
through a chute into a bin below, it is better to take the sample 
in small intervals from the end of the chute. When the same 
kind of ore arrives in sacks, a small sample is taken from the 
dryer top of every sack and a larger one from the centre, or 
the sack-sample is taken at once after emptying from what seems 
to be an average part. This last way sounds rather vague, but 
a person accustomed to emptying sacks will be able to take the 
correct sample. If the ore is not of uniform size, it has to be 
crushed in order to get a correct assay-sample. In that case, the 
moisture is best taken at the discharge of the crusher. When the 
crushing cannot immediately follow the unloading, and the ore has 
to be temporarily stored in a bin to be sampled down later on, the 
lumps may either be broken up and sampled with the fine part, or 
screened out and sampled separately, taking a correct proportion of 
coarse and fine in making up the final moisture-sample. 

In whatever way the sample may be taken, it is put into a deep 
tin vessel with a closely-fitting lid, so as to prevent the first part 
from drying while the other is being taken. The moisture-sample 
must be taken at once to the assay office. A certain amount, vary- 
ing from one to two pounds, is then weighed out twice for two 
moisture-determinations. These can be made when convenient, 
the weighed sample being transferred to a paper bag to await its 
turn. 

Before the moisture is determined the weighed sample is 
reduced to a uniform small size. This is done on the grinding- 
plate, the ore being crushed there to the size of a pea. If it is too 
moist to be crushed it is first dried somewhat. The final drying is 
performed in a shallow tin or granite-ware pan on an oil-stove, a 
sand-bath, on top of the boiler, or by any means that permits the 
controlling of the temperature. It is necessary to stir the sample 
at intervals with a spatula or spoon so as to dry it completely. A 
cool beaker-cover held a moment over the drying sample will show 
T)y the absence or presence of condensed vapor whether the drying 
is finished or not. 

The following form of receiving-book for ores shows the entries 
that are necessary to give the requisite information : 



SAMPLING, ASSAYING AND PURCHASING. 



47 



u 
o 

>< 




Shipper's 


«35 

y 

^3 


+3" 

'v -A 

& 

E 
O 




u 

EH 


^3 
bo- 


3 (1) 


be . 

fe O 
>s 


Placed 


Remarks. 


Name. 


Mark. 


9 ' 
m : 
c 6 


13 : 
PQ • 
5°' 





























§ 31. The Assay-Sample. — From a sample weighing a few 
ounces the metal value and the chemical composition of an ore are 
determined. The sample must therefore represent its average 
character. The most thorough way would be to crush the entire 
ore to a uniform small size, mix it, and then sample it down by one 
of the general methods to be discussed later on. If the ore is a 
sulphide, and has to be roasted before it is smelted, this prelim- 
inary fine crushing is necessary ; if, however, as is generally the 
case, the ore is treated raw, this crushing not only entails in- 
creased cost, but it is disadvantageous, as fine ore disturbs the 
uniform running of the blast furnace, increases the amount of flue 
dust, and retards the smelting. The best ore for the blast furnace 
is between the size of a man's fist and that of an egg ; it will be 
desirable to reduce it as little as possible below this. When, how- 
ever, the ore is rich and the metal-bearing part is irregularly dis- 
tributed in the gangue, as is mostly the case, it is necessary to 
crush finer before reducing to a smaller quantity. In sampling, the 
sample will grow finer as the bulk is reduced, this occurring more 
rapidly with high-grade than with low-grade ore. Reed 1 has, by 
an interesting calculation, arrived at certain definite maximum 
sizes which ores of a varying tenor in silver and gold may have 
when they are reduced from a hundred-ton lot down to the assay 
sample. His tabulated results, slightly condensed, are subjoined. 



Quantity of ore. Reducing 


Value of Silver in Ounces per Ton. 


Size of Ore. 


Highest : 300. 
Average : 50. 


Highest: 3000 
Average: 75. 


Highest : 

10,000. 

Average: 500. 


100 tons to 10 tons 


Cocoanut . 

Orange . . . 

Walnut... 

Pea. 

20 mesh. . . 


Fist 

E&g" 

Chestnut . 
Wheat. . . . 
25-mesh . . 


Fist 

Walnut... 
Chestnut. . 
Wheat. . . . 
50-mesh . . . 


Maximum 

permissible 

size of ore 

for given 

grade. 


10 tons to 1 ton 


1 ton to 200 pounds 

200 pounds to 5 pounds. . . 
5 pounds to bottle-sample. 



1 Op. cit., vol. vi., p. 357. 



48 METALLURGY OF LEAD. 

The breaking to walnut size is done in a crusher, from this 
down to the size of wheat by crushing in rolls ; smaller sizfcs are 
obtained by grinding in a mill or on a plate. 

While in practice these quantities and sizes are not rigidly 
adhered to, they are at least approximated. For example : it is 
not desirable to make one sample of a lot of moderately rich ore 
weighing over fifty tons. A hundred-ton shipment would be 
preferably divided into two lots and sampled separately, or be 
sampled by the carload (14 tons). Uniform low-grade ores can be 
sampled in 200 and 250 ton lots. With very rich ores the entire 
lot is taken as a sample. An ore must run very much lower than 50 
ounces per ton ; if the seller will be satisfied with a cocoanut size, 
he will insist on at least fist size, and prefer egg to walnut size. 
A 50-mesh sieve is generally considered to be too coarse for the 
final sample that is filled into the bottles ; the 80-mesh and often 
the 100 and the 120-mesh sieves, especially the 80-mesh, are in 
common use. 

Before sampling carbonate ores that are dry, they should be 
moistened with a hose, so that the fine dust, which often represents 
the richest part of the ore, may adhere to the lumps. If this is 
not done, it will be partly lost by being blown away when the ore 
is handled, or it may drop through the ore-heap to the floor and be 
partly lost there. In any case, the dust is liable to be unevenly dis- 
tributed, and thus errors may occur. For instance, a lot of ore 
sampled wet gave 400 ounces silver to the ton, sampled dry only 
390 ounces. 

Sampling is done either by hand or machine. There are four 
ways of doing it by hand : by quartering, by fractional selection, 
with the split shovel and by channelling. Machine-sampling is of 
two kinds : that which divides the stream of ore into two unequal 
parts and takes the smaller one continuously off as the sample, and 
that which takes intermittently at certain intervals the whole of 
the stream as sample. 

a. Quartering. — By this method the ore is reduced one-half in 
quantity at a time. It presupposes a thorough mixing of the ore, 
which is done by coning. The ore coming from the chute of a 
crusher or of rolls is wheeled away and dumped in a circle large 
enough for the two samplers to be able to stand inside. They work 
as partners, walking around the circle while forming the cone, 
remaining always diametrically opposite each other while they pile 
up the ore. In doing this, care must be taken that every shovelful 



SAMPLING, ASSAYING AND PURCHASING. 



49 



is thrown directly upon the top of the forming cone (Figure 1), bo 
that it may run and spread evenly down the sides. With experi- 




Quartering. 



enced men the circumference of this cone will he a true circle. 
When the ore is exhausted the floor is swept carefully and the 



50 METALLURGY OF LEAD. 

sweepings deposited on the top of the cone — not swept simply 
towards it — forming a ring of fine ore. A 50-ton cone is from 10 
to 12 feet high. The shovels used are long-handled and round- 
pointed. The ore being thus mixed, the cone is pulled down, the 
men working in opposite pairs (Figures 2 and 3). They begin 
near the top of the cone, and, walking round it, work it down from 
centre to periphery until it becomes gradually transformed into a 
truncated cone from 6 to 18 feet in diameter and from 6 to 12 
inches in height (Figure 4). A short-handled, square-pointed 
shovel is desirable for this work. With the sharpened edge of a 
straight lath two diametrical lines are drawn across the cone at 
right angles to each other. The ore is thus divided into four equal 
parts (hence the name quartering). Two opposite sections are 
removed to the bin (Figure 5), while from the other two a new 
cone is formed (Figure 6), and the process of quartering repeated. 
When the pile has become so small in quantity that it has to be 




BRUNTON'S QUARTERING SHOVEL. 

•crushed (see Reed's table) to attain a uniform sample, this is done 
by crusher or by rolls. The whole process is repeated until the 
sample is reduced to 100 or 50 or less pounds, forming the finish- 
ing sample, which is treated separately. Care should be taken to 
avoid losing any dusty part of the ore or working other material 
into the sample, the danger of error increasing as the sample 
becomes smaller. 

Quartering was formerly used as a starting method for the 
entire mass of the ore as it came to the smelter ; it has generally 
become a finishing method for reducing a part only of the ore. It 
is always very satisfactory. 

Two men will quarter a one-ton sample down to a few pounds 
in about two hours. Brunton ' has patented 2 a quartering shovel 
by whose use the operation is very much shortenened. It is (Fig- 

1 Engineering and Mining Journal, 1891, June 20. 
8 Patent No. 454,120, June 16, 1891. 



SAMPLING, ASSAYING AND PURCHASING. 51 

tire 1) a flat-bottomed steel shovel, 10 inches wide, turned up at 
the sides and divided into three compartments. The centre com- 
partment is closed at the back and has the width of one-quarter of 
the shovel. Thus, when the shovel has been pushed into a heap of 
finely crushed ore and filled, it is raised, and a "sharp rotary 
motion to the right " will discharge the ore from the outside com- 
partments to one side, forming the rejected ore pile, and leave the 
sample in the central compartment to be emptied on the other side. 
In a speed test a sample of one ton of ore was cut down with this 
shovel to 100 pounds in 15 minutes by one man. 1 

b. Fractional Selection. — This is a starting method. It con- 
sists in taking, while unloading, every fourth, fifth or tenth shovel- 
ful of ore for the first sample. This is then cut down in the 
same way. If the shovelling is done every time from the floor, as 
it should be, the ore without previous mixing will give a correct 
sample. 

The same method is used with sacks in making the first sample 
if the ore is known to be uniform in character, as otherwise the 
chances of error are too great. 

As in quartering, when the sample becomes small it has to be 
crushed. Fractional selection is rarely, if ever, used for finishing 
a sample. 

c. Split-shovel. — The tool used resembles a fork with a long 
handle, the prongs being replaced by from four to six troughs 
made of sheet iron. The width of the troughs, which is equal to 
the distance between them, depends on the size of the ore to be 
treated, and the largest piece of ore in the heap ought not to exceed 
one-fourth the width of the trough. Thus if the ore is half an 
inch, a 2-inch trough is required. The depth of a trough is such 
that a piece of ore striking the bottom will not rebound and 
fly out — say from 2 to 4 inches deep for a 2-inch trough. The 
length of the trough varies from 15 to 18 inches; the handle is 
that of a long-handled shovel. In sampling, the split-shovel is 
placed on the floor by one man, while the other, the sampler, facing 
the ends of the troughs, delivers the ore from ;i square-pointed 
shovel in ;i thin stream in the direction of the troughs. One-half 
of the ore is caught in the troughs, one-half passes into the spaces 
between. Re-sampling over the fork is continued until the troughs 

1 The writer is informed by disinterested parties that the rrsults with 
the Brunton shovel check well with their own results, obtained from 
regular quartet big. 



5 2 MET A LL URGY OF LEA D. 

are full, when it is lifted out and the contents discharged on a sep- 
arate heap, forming the reduced sample. The whole heap is passed 
over the fork in this manner, and the first reduced sample cut 
down similarly until it is necessary to crush finer, when it is passed 
over another split-shovel with smaller-sized troughs — say from J 
to 1 inch wide. 

That sampling with this tool requires even less previous mix- 
ing of the ore than with fractional selection is clear. It is also 
evident that sampling with the split-shovel can be used as a start- 
ing method only with pretty fine material. Smelters make the 
objection to the split-shovel that an undue amount of fine ore is 
liable to be caught by the troughs. This gives an incorrect sample 
if the fine ore assays differently from the coarse ore. The split- 
shovel is used at some sampling works for the entire sampling, but 
this is not common. It is rather the tool for finishing a sample, 
and is even used in the laboratory in a modified form as assayer 
riffle. 

This method works as fast as quartering, if not faster. 

d. Channelling. — This consists in spreading out the crushed 
and thoroughly mixed ore in a square, about 4 inches thick, and 
then taking the sample out in parallel grooves — say one foot apart — 
across the square, first one way, then the other. The sample is 
reduced by repeating the same operations. The greater the num- 
ber of the grooves (channels) drawn, the less thorough need be 
the preliminary mixing of the ore. 

This method, used formerly at Batopilas and Silver Islet, 1 is 
hardly used in any silver-lead smelting works now as a starting 
method. It is still in use as a finishing method, where the mixing 
of a small amount of fine ore is easily effected. 

e. Continuous Mechanical Sampling. — Quite a number of 
machines 2 for continuous sampling have been invented and are 
in use. The drawback to the method is, however, that a falling 
stream of ore is never uniformly mixed. In order to counterbal- 
ance this disadvantage, a preliminary fine crushing and sizing be- 
comes necessary to obtain a satisfactory sample. Why fine ore is 
undesirable for the blast-furnace has already been discussed. 

The Pipe Ore-Sampler, as represented in Figure 8 3 is an ap- 

1 Lowe : Engineering and Mining Journal, September 24, 1881. 

2 Egleston : Engineering, December 15, 1876. Reed : School of Mines 
Quarterly, ili., p. 253. 

8 Taken from a drawing of Messrs. Fraser and Chalmers, Chicago, 111. 



SAMPL1XG, ASSAYIXG AXD PURCHASING. 



53 



paratus of this class which is much used at present in lead-silver 
works. The finely crushed and sized ore is fed into the hopper 
a, ending in a small funnel b. Through this it passes over the 
divider c, which cuts it into two equal parts. One-half is dis- 
charged into the ore-bin d, while the other descends in the pipe 
and is thoroughly mixed by passing through the funnel b', and then 
being scattered towards the sides of the pipe before it is brought 




j> ; . 8.— PIPE ORE-SAMPLER. 
together again in a -mall stream by the funnel h\ which delivers 

ii t«» the Becond divider ''. The rejected on- falls into the bin '/, 

while the sample is collected in the sample-bin >. t<> be put again 
through the sampler nntij it has been sufficiently reduced in size. 
The ore-bin is closed by a wooden gate/] running in heavy iron 
guides. This is raised ami lowered with rack and pinion (g, //), 
the pinion -shafl i being turned by a hand-wheel j that i- keyed 

to it. The sample-bin 6 can be lucked, if desired. The main ,li s - 



54 



METALLURGY OF LEAD. 



advantage of the apparatus is that it is difficult to inspect and to 
clean. 

f. Intermittent Mechanical Sampling. — This method gives bet- 
ter results, as the entire falling stream of the ore is deflected at 
certain intervals to cut out the sample. The amount of ore taken 



Fig. 9. 



_V V VVVS_1 




out for the sample is regulated by the number of deflections that 
take place and the length of time they last. The advantage is 
that the bad effect of the irregular distribution in the stream of 
ore is neutralized, and thus fine crushing and screening made 
unnecessary. Rittinger x describes an apparatus of this kind. Two 
1 " Lehrbuch der AufbereituDgskunde," Berlin, 1867, p. 583. 



SAMPLING, ASSAYING AND PURCHASING. 



55 



modern machines by Brunton and Bridgman may serve as exam- 
ples : 

The principle embodied in the Brunton x sampler is shown by 
Figures 9 and 10. The ore coming down through spout C and 
the funnel B is divided by the deflecting chute A into two parts, 



ri 



Fig. 10. 




which arc discharged separately by the spouts J) and 77. The 
movement of the chute A is effected in the following way : to 
the arms of the pulley Q t driven by tin- belt J/, is attached a 
smaller wheel H, the fare of which is perforated by two rows of, 

say, twenty holes. Intotlie.se are inserted alternating! v on one side 
1 Trans. A. J. M. E. t xiii., p. 689. 



56 



METALLURGY OF LEAD. 



and the other a number of pins Z, and then fastened like bolts on 
the inside face by means of jam-nuts. When the wheel .Zf revolves, 
the pins will move the guides JV, iV alternately to the right and 
to the left, and these the driving-bar I, with the pitman J, which 
gives the deflecting chute a reciprocating motion. 

The proportion of ore that is taken out as a sample depends on 
the disposition of the pins X. If these are inserted half in the 
right-hand row of holes and half in the left, on opposite sides of 
the wheel-face, then the deflecting chute will be held to the right 
during one-half of a revolution of the wheel and to the left during 




the other half, and thus the ore-stream divided into two equal parts. 
If the division of pins is 10 ])er cent, to one side and 90 per cent, to 
the other, the deflecting chute will remain T 1 ¥ of a revolution on 
one side and take out ^ of the ore as a sample. It will thus be 
seen that by a suitable distribution of pins any percentage de- 
sired of a falling stream of ore can be recovered as a sample. The 
number of revolutions wheel XT makes per minute must be large if 
the sample is to be a correct one. The further reduction of the first 
sample can be effected, after crushing fine with rolls, by a smaller 
machine of similar construction, down to the finishing sample. 



Fig. 12. 




58 MET ALL UBQ Y OF LEAD. 

The Brunton machine gives satisfaction on account of its sim- 
plicity, the accuracy and speed of its operations, and its doing 
away with the necessity of fine-crushing the entire lot of ore that 
is being sampled. 

Figure 11 represents a cross-section through a sampling-mill 
using the Brunton sampler. The ore is discharged on the screen- 
plate #, and the coarse ore crushed in the Blake crusher b. Both 
tine ore and crushed coarse ore drop into the bin c, whence they are 
delivered by a belt-elevator c?at the top of the building to the fun- 
nel e. The deflecting chute/" divides the stream of ore into two 
parts ; the rejected ore passes down through the chute g, which 
discharges it into an ore-car to be transferred to a bed or bin. 
The sample is conducted by the chute h to the hopper i of the rolls 
j, and the finely-crushed sanij)le is then further reduced by quar- 
tering. 

Tlie Bridgman Sample?'. 1 — Type A, represented by Figures 
12, 13 and 14, approximately drawn to scale, may serve as an exam- 
ple of the machine. It occupies a floor-space 3 by 4 feet, and is 
7 feet 6 inches high. It consists of three hollow truncated cones, 
I., II., III. (called apportioners), driven by the pulley F, and three 
stationary concentric receptacles, /»*,, Tig and H. The first two dis- 
charge the original and duplicate samples they receive through the 
spouts T^ T 2 into the sample-boxes Z ly Z 2 ; the third discharges the 
rejected ore that falls into it through the chute S. 

Apportioners I. and III. (Figure 14) move in the same direction, 
making respectively 5 and 45 revolutions per minute ; apportioner 
II., moving in the opposite direction, makes 15 revolutions. Ap- 
portioner I. (Figure 14) consists of two concentrical rings having 
eight compartments, i x to X 8 . To each of these an adjustable spout 
is attached, one in the direction of J7 — 1, a second in that of Jil — 2, 
and six in that of JI — D. When the apportioner is in motion, 
M — 1 describes the circular path 1 — 1, J7— 2 the path 2 — 2, and the 
rest the path W— TK The intermediate (II.) and lowest (III.) ap- 
portioners have the same construction: a hollow truncated cone 
with the central discharge opening W t and four vertical chutes 
A^ — iVj and A" 2 — A^>, each of which represents one-eighth of the 
path covered by the spouts M — I and M— 2. 

If crushed ore is fed through the chute .Fit will fall equally into 
all the spouts ; thus, with every revolution of apportioner I., one- 

1 Trans. A. I. M. E., xx. 



Fig. ia 




■;*- B ^f^'^l^ 



CO METALLURGY OF LEAD. 

eighth of the ore passing through 31—1 goes to make up the origi- 
nal sample, another eighth passing through 31— 2, the duplicate 
sample, and the remaining six the rejected ore, which passes off 
through W, W, //(Figure 14) and spout S (Figure 13). The 
samples caught by 31—1 and 31- — 2 now undergo separately the 
same operations. To follow the one in 31— 1 : it is intercepted by 
apportioner II., which cuts out and delivers to D three-fourths of 
it, the remaining fourth passing through chute JV—1 to receive the 
same treatment by the apportioner III. Thus the amount a', cut 
out by 31—1 with every revolution of apportioner I. and forming 

a' 
one-eighth of the ore fed during the revolution, is reduced to — by 

apportioner II., and to ___ by the apportioner III., when it drops 
1(5 

into the sample-box Z x . Thus a 2 cut out by 31—2, undergoing the 

same operation, is reduced in like proportion and collected in box Z 2 . 

The machine differs from all others in taking out duplicate sam- 
ples and cutting them down as far as possible without re-crushing. 
It does its best work with ore that is not larger than 1-inch 
cube. It has a great capacity, varying from 15 to 25 tons per 
hour, is worked cheaply, and can be easily cleaned. 

Type B, described and illustrated in the reference given, is 
a smaller machine of similar construction furnishing only one 
sample. 

The general arrangement of a sampling-mill using the Bridgman 
sampler is shown in Figures 15 and 16, the scale being 8 feet to 1 
inch. The ore arriving in the box car F is unloaded after having 
been weighed on the i^latform scales X. At the same time it is 
sampled by fractional selection, the sample being delivered through 
chute S (located between the bins D) into car Q, which is dis- 
charged into hopper B l . From .this the sample passes through 
crusher C\ thence through hopper JP and over rolls B into the 
large sampler H (type A). The rejected ore drops into the bin of 
elevator E, which discharges into car 31. This brings the dis- 
carded portion to the bin D, where the main part of the lot is 
stored. Original and duplicate samples in boxes 1 and 2 are 
elevated by a chain-tackle to the tilting frame 2\ and discharged 
separately into hopper B 3 of the fine rolls 0, which deliver the 
fine ore to the small sampler (type B). This gives the finishing 
sample, weighing a few pounds, which now goes to the sample- 
grinder and thence to the assay office. 



Fig. 14. 




62 METALLURGY OF LEAD. 

g. Comjycirison of Methods, — With proper care an accurate 
Sim plecan be obtained by any of these methods. Fractional 
selection and mechanical sampling are the best for handling large 
quantities of ore ; small amounts are effectively sampled by quar- 
tering, fractional selection, and by the use of the split-shovel. The 
best finishing methods are quartering, using the split-shovel, and 
channelling. Finally, to compare hand sampling and machine 
sampling, it may be said that, while in the former you are de- 
pendent on the workman and the method is slow, expensive, and 
requires a large space to work in, machine sampling has not as 
yet displaced it, and a smelter will be best served by not depend- 
ing entirely upon either method, but using both as occasion may 
require. 

h. Finishing the Sample. — This is best done in a well-lighted 
room, the finishing-room adjoining the sampling floor. The sample 
must be protected from the wind, from dust, and from any 
possibility of being tampered with. The seller is permitted to 
watch the sampling through one of the windows, but not to enter 
the room, where the head sampler works alone. If the sample is 
moist, it is dried somewhat, so that it can be ground. Ore that is 
very moist or frozen when it comes to the works can generaHy be 
sampled down to one ton before it needs to be dried. 

The sample obtained by one of the different methods de- 
scribed, weighing from 50 to 100 pounds (the particles being, 
according to Reed's table, for 50-ounce ore not larger than a pea), 
has to be prepared for the laboratory. If larger it is ground either 
by hand or by machine to pass a 10-mesh sieve. If this is done by 
hand, a heavy muller, with head 4-J by 7 inches, is used on a circular 
cast-iron plate of from 3 to 4 feet in diameter and from 1 to 1^- 
inches in thickness. The plate, the working side of which is planed, 
stands free, so that the sampler can walk around it while grinding. 
Sometimes an oblong plate is used, 18 by 24 inches, with a rim J 
of an inch high on two or three sides and a muller head 4 by 6 
inches. More commonly the comminution is effected by a machine, 
the so-called sample-grinder, 1 which resembles a coffee-mill. The 
sample is now reduced to about ten pounds by quartering, using 
the assayer's riffle, rarely by channelling. In quartering a funnel is 
very handy to form the cone, which is then stirred down with the 

1 The makes of Fraser and Chalmers, in Chicago, and of Hendrie and 
Bolthoff, in Denver, are those generally used. 



Fig. 15. 
















1 




















































































































1 




















u 


-x 




i 








1 


c_ 
















64 MET ALL URGY OF LEAD. 

palm of the hand, with a 2 by |-inch lath from 12 to 18 inches long- 
held horizontally, or with a spoon held vertically. The assayer- 
rifile resembles a small split-shovel without a handle. The other 
tools are a small scoop and a flat paint-brush. 

The Bridgman laboratory sampler 1 (Figures 17 and 18) is a very 
handy machine to simplify the quartering, occupying a space 14 
inches square and being about 14 inches high. It has as its main 
parts a divider D (which is set in motion by hand, by clockwork, 
or by any convenient power), a funnel 1, and a receptacle 2, with 
discharges O x and 2 respectively. The ore filling the hopper F runs 
in a continuous stream on to the divider, which gives eight cuts to 
the revolution : four of these, forming the sample, are delivered 
into the funnel ; the other four, the discarded portion, are collected 
in the receptacle. 

From the 50 or 100 pounds either a single sample may be 
made, or two, called original and duplicate. In the latter case 
the two halves obtained of the ore are treated separately ; if the 
rime is used, that part which has been caught by the troughs is 
dumped alternately to one side and the other, forming two heaps, 
which are then worked up separately. The duplicate sample is 
useful in two ways : first, to verify the work ; secondly, to expose 
any tampering with the samples. When these are completely 
dried they are ground to pass a 20-mesh sieve, then well mixed 
by rolling on an oil-cloth or thin rubber cloth, and cut down to 
about 6ne pound. This is again ground on the plate to pass a 60- 
mesh (or more commonly an 80-mesh) sieve, and is then again well 
mixed by rolling on glazed paper, after which it is ready to be 
filled into the large-necked four-ounce sample bottles. 

To simplify this last operation Bridgman 2 has devised a mixer 
and divider, shown in Figures 19 and 20. The mixer, a large fun- 
nel with movable cover, is filled with the ore, well shaken, and then 
passed to and fro over the divider, which discharges the sample 
evenly into the four bottles beneath it. 

Sometimes scales of metal in the ore are too large to be ground 

with the rest to pass the sieve. They are then weighed, assayed 

separately, and the average result calculated by the following 

formula : 

S.a + P.a' 



S + P 



= a 



1 Trans. A. I. M. E., xx. 

2 Trans. A. I. M. E., xx. 



Fig. 17. 




Laboratory Sampler. 



66 



METALLURGY OF LEAD. 



S = weight of scales ; P = weight of pulp ; a and a', the respective 
assays in ounces per ton ; a", the average assay. The scales are 
best assayed all together, as dividing them is liable to cause 
error. 

Each sample is divided into three parts, filling three bottles : 
one for the seller, one for the smelter, and one for the control. 
Triplicate assays are made from each sample, and thus both 
sample and assay are checked. If much discrepancy is found 




MIXER AND DIVIDER. 



between the original and the duplicate, re-sampling becomes neces- 
sary ; if only little, it is averaged. 1 

It is hardly necessary to say that the sampling floor and all the 
apparatus must be thoroughly swept and cleaned after each opera- 
tion ; this is especially important in the finishing-room, as the 
smaller the sample becomes the more carefully must it be protected. 
The last traces from previously comminuted ore can be best re- 



1 "Leadville, Smelter's Regulations : 
rial, 1883, October 27. 



Engineering and Mining Jour- 



SAMPLING, ASSAYING AND PURCHASING. 67 

moved by using a small quantity of the new sample, taking care 
to return this to the general bulk. 

The charges made for sampling vary from $1.50 to $3 per ton, 
according to the grade and character of the ore. Rich silver ores, 
consisting mainly of gangue and silver mineral, and gold ores with 
free gold, often require triplicate samples, and always closer work 
than uniform low-grade ores. The last are sampled, if in large 
quantities, at $1 a ton ; and with concentrates requiring no crush- 
ing the price is often as low as 85 cents. 

§ 32. General Arrangement of Sampling Department. 
— The sampling department should be arranged so as to avoid all 
unnecessary handling of the ore. 

For hand sampling the simplest way for small works is to have 
the ore-bins below the main sampling floor, so that the rejected ore 
can fall at once into the bins and remain there until required for 
the blast-furnace. The distance must be sufficient for the ore to 
be discharged from the centre into any of the bins. The receiving 
floor need only be raised enough above the sampling floor that 
the ore discharged by the crusher shall fall into a wheelbarrow 
placed below, the mouth of the crusher being on a level with the 
receiving floor and near the middle of the edge. A distance of 4-J- 
feet between receiving and sampling floors is sufficient for a No. 5 
Blake crusher. The receiving floor is about a third as large as the 
sampling floor. The floors are made three inches thick and consist 
of two layers of planking, the upper being boards 1 inch thick, 
running across the heavier planks, and parallel with the mouth of 
the crusher to facilitate the shovelling. In front of the crusher the 
floor is protected by a sheet-iron plate, say ^ inch in thickness, 
fastened by countersunk screws. 

Another arrangement, found in larger works, is to place the 
sampling-room opposite the ore-bins on the other side of the track. 
This is laid so low that the floor of the car is on the same plane as 
the sampling-room and the run ways on top of the ore-bins. The 
sampling-room is a simple, oblong building with stalls on the four 
sides to keep the samples. It has only one floor, on which are 

placed crusher and rolls. The finishing-room is partitioned off 
from the main room. While the ore is being unloaded from the 

car, the sample is taken by fractional selection and wheeled to 
one of the stalls on the sampling floor, the bulk of the ore passing 
st raight to the bins. 

In machine sampling, which is used only in the largest works, 



68 



METALLURGY OF LEAD. 



the sampling mill is never directly connected with the smelter 
building. Its arrangement will depend on the character of the 
machine and the configuration of the ground. 

A lot of ore, after being discharged into a bin or on a bed, 
is spread out evenly before another lot is received. Each subse- 
quent lot will thus form a distinct layer over the preceding one. 

A special book is kept for entering the contents of the different 
bins and beds, the averages being calculated according to the 



formula in § 20. 



RECORD OF BINS AND BEDS. 



j 


B. 


1 


Date. 


Contains Lots 
No. 


Total Weignt, 
Tons. 


Sr.2 

S3 O 

- c 

a - 
>•_ 
< - 


Average 


No. 


1 

5 


1! 


I 

£ 

71 


Wet. 




Oz. per 
ton. 


Per cent. 


< 


$ 


X5 


3 

o 


c 




CO 





































Assay. 


Remarks. 


Per cent. 




Si0 2 . FeO. 


MnO. 1 CaO. 


MgO. 


BaO. 


A1 2 3 . 

























§ 33. Receiving and Sampling of Fluxes and Fuels.— 

Fluxes and fuels on arriving at the smelter are weighed on plat- 
form scales and unloaded as near the feed floor as possible. It is 
necessary to keej) large quantities of fluxes always on hand. As 
they take up a great deal of room and are not injured by the 
weather, they are generally left outside the building. The fuels 
are placed in sheds for protection. The breaking of fluxes is done 
by hand or in a coarse-set crusher. The latter method makes more 
fines. The most desirable size for iron ore and limestone is that 
of a man's fist ; for quartz, that of an egg. 



SAMPLING, ASSAYING AND PURCHASING. 69 

In sampling fluxes and fuel, a so-called grab-sample gives suffi- 
ciently near results. This consists in taking out bits or handfuls 
here and there from the top, middle, and bottom of the heap until 
about 50 pounds of iron ore, 30 of limestone, or 10 of coke have 
been taken. From the grab-sample the moisture-sample is taken, 
and the remainder crushed fine and cut down to be analyzed in the 
same manner as a regular ore-sample. 

§ 34. The Assaying of Lead-Silver Ores.— This is for two 

purposes : first, to ascertain the amount of metal contained in the 
ore ; second, to determine its chemical composition so far as this 
relates to the making up of the smelting charge. 

Assays are made in the dry way to ascertain the amount of 
lead, silver, and gold an ore contains ; all the other elements of 
importance are determined in the wet way. 

For the dry assay two kinds of furnaces are required, the cruci- 
ble and the muffle furnace, the two being usually placed side by 
side. Drawings of furnaces are given in all books on assaying, 
and are unnecessary here. 1 A few comments may, however, be in 
place. With crucible furnaces the grate should be of wrought -iron 
bars that can be shaken and drawn out from a horizontal slit above 
the ash-pit. Portable muffle furnaces are used in small works and 
in places where bituminous coal is not obtainable. One of the best 
models in the market is " Brown's portable assay furnace." In 
large works permanent furnaces buiit of common brick and lined 
with fire-brick are the rule. The use of bituminous coal is prefer- 
able to that of carbonized fuel or anthracite, as it is easier to reg- 
ulate the temperature and the bottom of the muffle remains hotter. 
In building such a furnace special attention is to be given to hav- 
ing the passages of the correct width, if the fuel is to surround the 
muffle, or to giving the inside wall of the furnace the form of the 
muffle, if the flame is to play within it. With coke, anthracite or 
charcoal the passages ought to be about 4 inches wide at the sides, 

1 W. L. Brown, "Manual of Assaying," Chicago, 1889; P. de P. 
Ricketts, "Notes on Assaying," New York, 1879; C. Balling, "Probir- 
kunde," Brunswick, 1879; "Supplement," Berlin, 1887; B. Kerl, "Metal- 
lurgische Probirkunde," Leipsic, 1882; "Supplement," ibid,, 1887; B. 
Kerl, " The Assayer's Manual" (an abridged treatise of the Larger work), 
transl. by W. T. Brannt and W. H. Wahl ; new edition by F. L. Garrison, 
Philadelphia, 1890; A. Biche, "L'art de I'essayeur," Paris, lsss. Bee 
also Wood, School of Mines (Quarterly, xii, p. 18G ; Engineering and Mining 
Journal, 1891, September 26. 



TO METALLURGY OF LEAD. 

so that the fuel need not be broken into small pieces, and 6 inches 
at the back, as this part chokes up readily. With bituminous 
coal the flue at the back can have the same width as that at the 
sides, about 2 inches. The vertical distance from the grate bars to 
the bottom of the muffle must be at least one-half the length of the 
muffle. Muffle furnaces are better fired from the back than from 
the front. The muffles used at large smelting works hold from 12 
to 16 2^-inch scorifiers. 

Lead assays are made both in muffle furnaces and in crucible 
furnaces. With the former the models of the Denver Fire Clay 
Company are commonly used : the 5-gramme (2} inches high by 
21 inches w T ide at the top) and 10-gramme crucibles (3 inches high 
by 2J inches wide at the top), called by the Battersea Company 
"Colorado A A " and "Colorado A." In the crucible furnace, 
Hessian and Battersea crucibles, 4-J inches high and 3 inches wide 
at the top, form a convenient size for 10 grammes of ore. The 
muffle assay is generally preferred to the crucible furnace assay 
at smelting works. Already, in 1870, Percy * proved that a wrought - 
iron crucible was preferable to a clay crucible, and this has been 
substantiated over and over again since. The reasons are the 
shorter time required for the assay and the more intimate contact 
of the iron with the ore. 

Every laboratory contains two or more mixtures for assaying 
lead ores. They are measured into the crucible first, filling it 
about two-thirds, and the ore is added afterwards. The table on 
the opposite page gives a number of charges for special kinds of 
ores and for slag, in detail, and a few mixtures for general use. 

Silver assays are uniformly made by scorification, as it is 
adapted to all kinds of ores, and gives excellent results 2 unless 
organic matter 3 is present. In that case the crucible assay is pref- 
erable. 

In scorifying, 0.1 assay ton of ore is mixed with 1 ton of gran- 
ulated lead and 1 gramme of glass borax added. A 2^-inch scorifier 
is used. Results of triplicate assays should check to within 0.5 of 
an ounce per ton. If the ore is refractory, or if it contains volatile 
metals, half the usual quantity of ore must be used in the 2^-inch 

1 " Lead," p. 104. 

2 Berg- u. Hiitten-mdnninche Zeitung, 1867, p. 85 (Arents) ; 1867, p. 
1C2 (Fournet) ; 1874, p. 68 (Richter and Hiibner) ; 1877, p. 232 (Richter and 
Hiibner). 

3 Berg- u. Hiitten-mannische Zeitung, 1886, p. 441 (Gorz). 







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cr 
o c o 




p. 


81 * m 


! to m ; tc cr 


— ~ p 
S P 5| 




►T3 






a 




- 










<r» 


9* w : 


. io\*- N i- 1 to or 






Q 




s 




O 


<a> 


2> 






c 




• w ,-k • iiO 


CO 








3 




w 




* 5 t- , 
3 


: (-»-: ^s 


I 


i 




<x> 










* M 


: : m tc <c g 










|U 








►- 1 . . 


"* • C • • © 
* 


Slag. 





79 



METALLURGY OF LEAD. 



scorifier to leave room for the extra amount of lead and borax, the 
latter being added from time to time as the case may require. 

The following table for scorification assays is made up prin- 
cipally from Kerl's manual, already quoted : 



Weight of Argentiferous Substance 
0.1 assay ton. 



Galena, pure 

•' -with blende or pyrite 

"Wall accretions 

Lead matte, rich in iron 

" " copper or nickel bearing 



speise 
carbonate 



Blende 

Copper ores or matte. 



Grammes 
Test Lead. 



Grammes 
Glass 
Borax. 



Gray copper ores. 
Antimonial ores. , 
Arsenical ores. . . . 



Telluride ores , 

Dry silicious ores. . . 
Dry basic (iron) ores. 
Hearth accretions. . , 
Cupel-bottom 



15-18 
24-36 
30-40 
27-36 
36-48 



30-60 
10-15 

30-45 
36 

36-48 

48 

up to 48 

50 
35^0 
35-40 
2.5-50 
20-25 



up to 5 
0.6-0.9 
0.3-0.6 
0.5-1.0 

0.5-1.0 

0.5-0.7 
up to 0.5 

0.3-0.6 
0.3-0.5 

0.3-0.5 
up to 3.0 
up to 1.5 

0.3 

up to 0.5 

0.3-1.0 

up to 1.5 

up to 0.75 



Low temperature with 
copper; the button may 
have to be rescorified. * 

High temperature; the 
button may have to be 
rescorified. 



Remarks. 



High temperature. 
Low temperature : 

button may 

rescorified! 
And more borax. 



the 
have to be 



High temperature: often 
addition of litharge. 



Add litharge ; 
the button. 



rescorify 



If necessary, first evap- 
orate to drvness with 
HXO,. 



The gold of an ore is determined by dissolving the silver but- 
tons obtained from triplicate scorification assays. 

The scorification assay is not suitable for slags, because they 
contain too little silver, viz. : from 1 to 1\ ounces per ton. With 
a crucible assay, taking 1 assay ton of slag, a silver button of satis- 
factory size is obtained. The usual way is to prepare the smelting 
mixture in considerable quantities beforehand. For the assay, it is 
measured out to fill two-thirds of the crucible. From one-half to 
two-thirds of the amount is mixed with the slag, the rest is added 
on top, and this again covered with salt. A nail is then inserted to 
assist in liberating the lead. 





Crucible 


Mixtures. 








Litharge 

Sodium bicarbonate.. 




... .parts. 


1 

0.5 

0.*25 
0.3 


1 
2 

0.5 
0.3 


1 

1 

1 

0.5 

0.2 


Potassium carbonate. 

Borax, dried 

Argol 






Flour 




it 









SAMPLING, ASSAYING AND PURCHASING. 73 

The wet assays of other components of the ore, as well as of the 
fluxes, will be treated under the heads of matte (§ 79) and slag (§81)- 

§ 35. Purchasing of Lead-Silver Ores. 1 — In purchasing 
lead-silver ores the character of the lead-bearing mineral and the 
chemical composition of the gangue have to be considered. If the 
lead mineral is a sulphide, the ore has generally to be roasted ; this 
is not necessary if the lead is present in the f orm of carbonate. Fur- 
ther, an ore may be either self -fluxing, acid, or basic, i.&, requiring 
no fluxes, requiring a base (iron, manganese, lime), or requiring an 
acid (silica) to form a desirable slag. Lead-silver ores are com- 
monly acid, thus the basic ores command a higher price. In 
purchasing basic ores, the base paid for is iron, with its substi- 
tute, manganese. The so-called "base excess" is that amount of 
-available iron and manganese which is obtained by adding the per- 
centage of metallic iron to that of metallic manganese, and deduct- 
ing the percentage of silica. The impurities in an ore affect 
its price. They may necessitate a preliminary roasting (sulphur, 
arsenic), may impair the fusibility of the slag (zinc, magnesia, 
baryta), or may cause loss of lead and silver by slagging or vol- 
atilization (zinc, arsenic, antimony), or finally may render the lead 
impure (zinc, arsenic, antimony, copper). The price paid for an 
ore will therefore be in inverse ratio to the percentage of impuri- 
ties present. 

A smelter, especially if so located as to draw its supplies from 
a number of mining districs, treats not only argentiferous lead 
ores, but extracts the precious metal also from real silver ores, 
called dry ores on account of their want of lead, by mixing them 
with ores that contain more lead than is required for the charge. 
Smelting can thus often compete with milling on account of the 
higher percentage of precious metal extracted from the ore. 

In bidding for an ore deductions are made for loss and for cost 
of smelting, which will vary from the causes just mentioned. In 
the smelting charges are included the cost of shipping and refining 
the base bullion obtained. 

The value of lead is given in units of 20 pounds to the ton of 2,000 
pounds avoirdupois. Its price is regulated not simply by the fluc- 
tuations <>f tlif market, but also largely by the scarcity or abun- 
dance of available lead ore at the works. If the lead runs Lower 
than 5 per cent, it is not paid for. 

1 KirchhofF, " United States Geol. Survey: Mineral Resources of the 
United States," 1885, p. 251. 



74 MET ALL URGY OF LEAD. 

The value of silver and gold is given in ounces per ton. The 
price of silver is regulated by the New York quotation of the 
day, that of the gold varies slightly according to the amount 
present, which is generally small. The gold in an ore is not 
counted if it runs less than 0.1 ounce per ton. 

Competition has forced works to pay something for copper con- 
tained in sulphurets, if it runs as high as 5 per cent. 

There are various ways of arriving at the minimum smelting 
charge for an ore that is offered for purchase. The following 
gives very satisfactory results, and can be made to suit all cases. 
It is based on the cost of smelting what is called a neutral ore, and 
then debiting and crediting the actual ore for which the smelting- 
charge is to be fixed, as it differs in composition from the standard. 
A neutral ore is one in which the insoluble residue is equal to the 
sum of the iron (Fe) and the lime (CaO) it contains, and in which 
the lead amounts to 13 per cent. Taking Denver and Pueblo rates, 
f.o.b. works, 1 the manner of figuring may be shown by two examples: 
1. A sulphide ore containing : 

Ag. Pb. Au. Si0 2 . Fe. CaO. Zn. 

75 ozs. 13 %. 0.5 ozs. 25 %. 35 % 10 %. 

Debit : 

Smelting $4.50 

Roasting 2.50 

Fluxing 0. 00 (there is an iron excess and enough lead). 

Zinc charge 5.00 (10 units @ 50 cents). 



$12.00 
Credit : 

Gold $0.50 ($20.00 an ounce received by the refiner 

and $19.00 an ounce paid to the miner). 
Iron excess 1 .00 (35 -25 = 10 units @ 10 cents). 



$1.50 
The minimum charge to be made for treating this ore will be 
812.00 — $1.50 = §10.50 per ton. 

1 Cost of smelting, per ton $4.50 

" roasting, " 2.50 

Pay for gold , per ounce 19. 00 

" lead, " unit 0.10 

Charge for zinc, per unit 0.50 

" " grade silver (every 50 ounces silver above 100 

ounces per ton) 1 . 00 

Pay for iron or lime excess, per unit 0.10 

Charge for silica excess, " " 0.10 



SAMPLING, PURCHASING AND ASSAYING. 



75 



The value of the ore is calculated on the following basis : 

Lead: 30 cents a unit is paid when the price of lead in New 
York is $4 for 100 pounds, which is called paying on a four-dollar 
basis. The pay grows 1 cent for every rise of 5 cents for 100 
pounds in the New York market. 

Silver: 95 per cent, of the New York quotation for the day. 

Gold: $19 an ounce. 

The value of the ore for which the smelting charge is calcu- 
lated will be per ton : 

Lead 13 units @ 30 cents =$3.90 

Silver 75 ounces @ 95 cents = 71 .25 

Gold 0.5 ounce @ $19.00 = 9.50 



$84.65 
Smelting charge 10 . 50 



Net value of ore per ton 
2. A dry siliceous ore containing: 
Au. SiO,. 



$74.15 



Pb. 
4*. 



60 



Fe. 

10*. 



CaO. 
12 #. 



Zn. 



Ag. 
250 ozs. 
Debit : 

Smelting $4.50 

Roasting 0.00 

Fluxing . 90 (13 - 4 = 9 units lead @ 10 cents). 

Silica excess 3.80 (60 -22 = 38 units silica @ 10 cents). 

Zinc charge .... 3.00 (6 units @ 50 cents). 

Grade silver 3.00 (250-100 = 150 ounces ; $1.00 for every 

50 ounces.) 



Credit : none. 



$15.20 



The smelting charge will be $15.20 per ton. 
The value of the ore per ton is : 

Lead 4 units, no pay = $ 0.00 

Silver 250 ounces (u) 95 cents = 237.50 

Gold none = 0.00 

$237.50 
Smelting charge 15.20 



Net value of ore per ton = $222.30 

§ 36. Purchasing Non-Argentiferous Ores. 1 — In the Mis 

sissippi Valley this is not done by bidding in the market. The 



1 Clerc: Engineering and Mining Journal, July 4, 1885, p, 4. 



76 METALLURGY OF LEAD. 

" buyer" goes to the different mines and, guided by previous experi- 
ence in purchasing from the same mine, offers a certain amount for 
the concentrated galena. The price paid for pure concentrated 
galena is about $48 per ton. 

§ 3V. Purchasing of Fluxes and Fuels.— In regard to the 

purchase of fluxes and fuels there are no general standards as there 
are with ores. The manner of buying and price paid vary accord- 
ing to local conditions. 



PART II. 



THE METALLURGICAL TREATMENT 
OF LEAD ORES. 



' 



PAET II. 
THE METALLURGICAL TREATMENT OF LEAD ORES. 1 

§ 38. Classification Of Methods.— Lead ores are treated ex- 
clusively in the dry way. If free from silver, the resulting lead 
goes to the market, after it has been purified by liquating and 
poling. If the ore is argentiferous, the silver passes for the most 
part into the lead (base bullion), which has then to be desilverized 
(Part III.). Wet methods 2 and the electrolytic 3 extraction have 
been tried, but so far without success. The smelting of lead ores 
is carried on in furnaces of various forms. They may be classified 
under three heads : 

The Reverberatory Furnace. 

The Hearth Furnace. 

The Shaft Furnace. 

1 Percy, ''Metallurgy of Lead," London, 1870.— Phillips-Bauerman, 
"Elements of Metallurgy," London, Philadelphia, 1887, pp. 564-658.— 
Balling, " Metallhuttenkunde," Berlin, 1885, pp. 49-165.— Kerl, " Metall- 
htittenkunde," Leipsic, 1881, pp. 1-128.— Stolzel, " Die metallurgie," Bruns- 
wick, 1863-86, pp. 851-1050.— Rivot, "Traite de metallargie," vol. ii., 
Paris, 1872 — Griiner, " Sur l'etat actuel de la Metallurgie du plomb," 
Ann. d. Mines, 1868, xiii , p. 325 ; also as pamphlet : Dunod, Paris. — Griiner, 
" Notes additionelles sur l'etat actuel de la metallurgie du plomb," Ann. 
d. Mines, 1869, xv., p. 519; also as pamphlet : Dunod, Paris.— Cahen, "La 
metallurgie du plomb," Paris and Liege, 1863. 

2 Berg- u Huttenm. Ztg., 1880, p. 1. (Schaffner), p. 2. (Maxwell).— Wag- 
ner's Jahresberichte, 1877, p. 151 (Meunier). 

3 Berg- u Huttenm. Ztg., 1383, p. 252 (Kiliani), p. 378 (Blast-Miest). 



CHAPTEE VI. 
SMELTING IN THE REVERBERATORY FURNACE. 

§ 39. Introductory Remarks. — The process carried on in 
the reverberatory furnace is the Roasting and Reaction Process 
(§ 9). The Precipitation Process (§ 7) was formerly used with 
raw sulphide ores at Vienne 1 (France) and at Chicago, 2 and with 
roasted ores at Par and Point 3 (Cornwall), but has now become 
obsolete on account of the high cost and loss of metal. 

The roasting and reaction process in the reverberatory furnace 
has the following advantages : the ore is treated in the raw state, 
the apparatus is inexpensive, inferior raw fuel is used, hardly any 
fluxes are required, the bulk of the metal in a pure state is quickly 
extracted at a low temperature with little loss by volatilization, 
and if the ore is argentiferous, the larger part of the silver follows 
the lead, and only a small quantity is left in the residue, which is 
either thrown away or treated at a higher temperature in the blast- 
furnace. The great disadvantage of the method is that it is very 
much limited by the character of the ore. To be suited for the 
reverberatory furnace an ore must be a rich galena or a mix- 
ture of galena with carbonate (the former prevailing), that does not 
contain less than 58 per cent, of lead. It may not contain over from 
4 to 5 per cent, silica ; and the non-siliceous associated minerals, 
such as blende, pyrite, chalcopyrite, calcspar, barite, may be present 
only in small quantities. The process requires much fuel and 
many hours of skilled labor per ton of ore treated. 

It consists of two operations, one following closely upon the 
other, and both being repeated several times. 

1. Oxidation. — The ore, crushed fine enough to pass a 4- or 
5-mesh sieve, is spread in a thin layer over the hearth of the fur- 
nace and is heated gradually to a low red heat. The roasting is 
•carried on in such a way that only a part of the lead sulphide is 

1 Kerl, "Metallhuttenkunde," p. 24. 

8 Trans. A. I. M. E., ii., p. 279 (Jerneg-an). 

8 Percy, "Lead," p. 257. 



80 METALLURGY OF LEAD. 

converted into oxide and sulphate, the rest remaining undecomposed. 
The temperature at which this roasting is carried on and the time 
given to it depend on the character of the ore. Pure galena re- 
quires a low temperature to avoid agglommeration ; if pyrite, 
blende or calcite are present the roasting can be accelerated ; the 
lower the temperature, the more sulphate will be formed. During 
the operation the fire on the grate is kept low and open and the 
charge is raked frequently, to expose as much of the ore as possi- 
ble to the action of air and heat and to prevent agglomerating, 
which obstructs oxidation. 

2. deduction. — The second operation is that of raising the 
temperature, so that the oxygen compounds may react on un- 
changed sulphide. The resulting lead runs down the inclined 
hearth and collects in a basin, the sulphurous acid passes off 
through the flue, and the residue remains on the hearth. The tem- 
perature during the reduction period must be low, as the reactions 
do not take place if the ore is melted. By filling the grate well 
up with fuel the required temperature is obtained and unconsumed 
air excluded. The charge is stirred at intervals to bring sulphide 
and oxide constituents into intimate contact. 

As it. is not possible to roast a large amount of lead ore uni- 
formly in one operation, the first reaction that takes place on rais- 
ing the temperature will not extract all the lead. The resulting 
pasty residue will be rich, consisting mainly of lead sulphide with 
some oxide, sulphate, silicate, and gangue. The temperature is 
lowered and air is admitted. A second roasting takes place and 
is followed by a second reacting. It takes several rej^etitions of 
the process to extract the bulk of the lead. With each one the 
temperature must be slightly raised as the amount of lead dimin- 
ishes. To counteract the melting of the charge, slacked lime is 
added, which acts mechanically by rendering the charge less fusi- 
ble and more spongy. It also assists the process chemically by 
liberating the lead (§ 7) and by decomposing the sulphide, thus 
helping the silver in the residue to pass into the lead. Towards 
the end of the process there will not be enough lead sulphide left 
to react on the excess of lead sulphate (§ 9) and lead oxide. To 
reduce these to sulphide and metal and to make the charge more 
porous, charcoal or coal is mixed into it. Then the roasting and re- 
acting can proceed again. Each successive operation will be of 
shorter duration than the preceding one and the lead each time a 
little less rich in silver. The first lead may contain four times as 



SMELTING IN THE REVERBERATORY FURNACE. 81 

much silver as the last. 1 The products of smelting in the rever- 
beratory furnace are : 

1. Lead, holding in suspension particles of ore and other solid 
matter, which are removed by liquating and poling (§ 109). If 
the ore contains arsenic, antimony, or copper, some of these ele- 
ments combine with the lead and have to be removed by refining 

(§ its). 

2. Gray Slag, a more or less matted mixture of lead, lead sul- 
phide, oxide, sulphate, silicate, gangue, cinders, and lime in vary- 
ing proportions. Its tenor in lead and silver depends on the char- 
acter of the ore and on the degree to which the residue has been 
treated in the furnace. In some cases it is crushed and washed to 
save only the metallic lead ; in others, especially with silver-bear- 
ing ores, it is smelted in the blast furnace. 

3. lluedust is composed of particles of unchanged or oxidized 
ore, volatized lead that has been converted into oxide, carbonate 
and sulphate, and parts of fuel. If the ore contains blende, oxi- 
dized zinc compounds will also be found. The amount of Hue- 
dust formed will vary with the temperature at which roasting and 
reacting operations have been carried on and also with the skill of 
the furnace-man in manipulating the furnace and the charge. As 
it consists principally of oxidized compounds, it is worked in with 
subsequent charges and shortens the time required for roasting. If 
very impure, e.g., from arsenic and antimony, it is smelted in the 
blast-furnace (?j 52) witli the gray slag. The resulting lead is 
hard and has to be refined (§ 105). 

4. Hearth Bottom, consisting of hearth material soaked to 
some depth with metal. It is worked up in the same manner as the 
residues. 

S 40. Influence of Foreign Matter.— The quantity and 

quality of lead that can be obtained from a given lead ore will de- 
pend largely on the nature and proportion of the other constitu- 
ents. These may be silica and argillaceous matter, oxides of iron, 
lime-tone (dolomite), barite, fluorspar, pyrite, chalcopyrite, blende* 
antimony, arsenic, silver (gold), and oxide lead ores. 

Silica and Argillaceous Matter have an injurious influence in 
both stages of the process on account of their readiness to com- 
bine with lead oxide (§ 6). It has been found by experiment that 

1 Ben/- u. inittenm. Ztg., 18G0, p. 359; 18C3, p. 886 (Fallize) : 1871, p. 
152(Bouhy). Zeitschrift fur Berg-, UiUtci- u. Salinen-Wen », siv., i>. 382 

(Teichmann). 



82 METALLURGY OF LEAD. 

with more than 5 per cent, of silica an ore cannot be treated by 
the roasting and reaction process. But even such a small amount 
as | per cent, makes itself felt by coating particles of ore with the 
silicate that has been formed, thus preventing the action of the air 
during the first period and obstructing the reactions when the tem- 
perature is raised. 

Oxides of Iron. — Siderite is sometimes found with galena ores, 
but most of it can be removed by dressing the ore before smelting. 
The small amounts which remain with the galena quickly lose 
their carbonic acid during roasting, and the resulting magnetic or 
ferric oxide acts as a stiffening ingredient while the lead is being 
extracted. 

Limestone (Dolomite) acts on the whole advantageously during 
the entire process, as it hinders the fusing of the charge. Any 
chemical action it may have is so slight that it can be regarded as 
practically inert matter. It loses some of its carbonic acid and is 
to a small extent converted into sulphate. Like all indifferent 
substances it will retard somewhat the roasting by preventing the 
air from having free access to the particles of galena and obstruct- 
ing the reactions by interfering with the necessary intimate contact 
of sulphide and oxide. The highest allowable amount is 12 per cent. 

Barite remains practically unchanged in the reverberatory. 

Fluorspar, the same. If fluorspar and barite are present to- 
gether, they may increase the fusibility of the charge by combin- 
ing with lead sulphate. 

Pyrite is beneficial in the first stage. It favors the forming of 
lead sulphate (§ 9) and assists the oxidation of galena through 
sulphur trioxide liberated by the decomposition of ferrous or fer- 
ric sulphate. Small quantities of pyrite have also a favorable 
effect during the reaction period, the ferric oxide making the 
charge less fusible. If present to a considerable amount, say 
from 10 to 12 per cent., 1 too much iron sulphide remains in the 
charge and is liable to form double compounds with the lead 
sulphide. With from 35 to 40 per cent, of pyrite 2 the reverbera- 
tory process has to stop. 

Chalcopyrite has no special effect during the roasting and be- 
haves in a manner similar to that of iron pyrite in the reaction 
period, with this addition, that some of the copper enters the lead 
and impairs the quality. 

1 Bouhy : Annates des Mines, 1870, xvii., p. 179. 

2 Bouhy, Op. cit., p. 178. 



SMELTING IN THE REVERBERATORY FURNACE. 83 

Blende. — During the roasting period blende is partly converted 
into oxide and sulphate, and the latter perhaps somewhat decom- 
posed, but much blende remains unaltered. From 4 to 5 per cent. 1 
assists the roasting ; 10 or 12 per cent. 1 prolongs it greatly and re- 
duces the output of lead; with from 20 to 24 per cent. 2 very little 
lead is obtained; and from 35 to 40 per cent., as with pyrite, stops 
the process. The loss in silver with blende-bearing ores is princi- 
pally a mechanical loss that takes place during the roasting period. 

Antimony. — This occurs with some galena ores as a si'mple or 
multiple sulphide. It has a deleterious effect even if present in 
small quantities of 2 or 3 per cent. It is readily fusible and causes 
caking of the ore. The sulphide and the oxide are both volatile, 
thus causing loss. The oxide also combines readily with lead ox- 
ide, which is retained to a great extent by the gray slag ; further 
sulphide and oxide of antimony react upon each other, giving 
metallic antimony, which is in part volatilized, causing again loss, 
and the remaining metal finally combines with the lead, making 
this hard. Thus antimony is probably the worst metal that can 
be associated with the lead. 

Arsenic. — The pyrite formed with lead ore sometimes contains 
arsenic. Next to antimony, arsenic is the worst enemy of the 
process. It causes losses : by volatilization, through the arsenious 
acid formed in roasting ; by slagging, through the combination of 
arsenic acid with lead oxide ; and finally by the reduction of both 
to metal. This combines with the lead and impairs its quality. 

Silver end Gold. — Most of the silver of galena ores passes off 
into the lead ; gold behaves in a similar way. 

Oxide Lead Ores, such as anglesite and cerussite, assist the 
operation, as they reduce the time of roasting. 

§ 41. Classification of Reverberatory Methods.— The prac- 
tice of the roasting and reaction process varies in different smelt- 
ing works. At, some the principal stress is laid on extracting 
as much lead as possible in the reverberatory, while others aim 
to obtain only the major pari of the lead in the reverberatory and 
to extract the rest from the rich residue by smelting it in the blast- 
furnace ; they thus recover a Larger percentage of lead. Then, 
some establishments roast the ore slowly at a low temperature, 
which is advisable foT the recovery of a large percentage of lead, 
while others hasten roasting by raising the temperature quickly to 

1 Bouhy, 0)>. ril.. p. 178. 

2 Itivot, " Metallurgie du plomb," p. 46. 



S4 METALLURGY OF LEAD. 

the permissible limit, the aim being to save labor and time, which 
is done at the expense of the lead. As regards the form' and size 
of furnaces and the position of the lead-well there are also char- 
acteristic differences. 

The reverberatory furnace practice may be classed under three 
main heads : 

The Garinthian Method. 

The English Method. 

The Silesian Method. 

§ 42. The Carillthian Method. — The characteristics of this 
method are the smallness of the charge, the slow roasting, so 
that for every part of lead sulphide one part of sulphate and at 
least two of oxide are formed, the low temperature at which all 
the operations are carried on, and the aim to extract all the lead 
in the reverberatory. The hearth is inclined towards the flue and 
the lead is collected outside of the furnace. 

A. Lead-Smelting at Haibl, 1 Carinthia. — The ore that is 
worked is galena (partly coarse, with from 72 to 75 per cent, of 
lead, and partly fine, with from 67 to 73 per cent, of lead); in 
exceptional cases the lead contents go as low as 58 per cent. The 
other constituents of the ore are anglesite, cerussite, wulfenite, 
blende, willemite, calcite, fluorite, and abestos. As seen from the 
assay these other constituents form only a very small quantity. The 
following analyses made by Phillips 2 in 1845 show the composition 
of low-grade coarse and fine ore : 



PbS... 
PbC0 3 . 
ZnS... 
Sb 2 S 3 . 
CaC0 3 



oarse. 

76.6 


Fine. 
76.0 


4.0 


4.0 


13.2 

0.2 


13.0 

0.2 


4.6 

0.4 


5.0 

0.4 



Insoluble 

99.0 fc8.6 

The drawings (Figures 21-24) of the furnace show an inclined 
hearth with only one working-door, g, below the flue d. On the same 
side is the door b leading to the fireplace. The grate, slightly more 
inclined than the hearth, is parallel to the long axis of the furnace. 
It is built of stone and has six transverse openings. The fire-bridge 

1 Thum : Berg- u. Hiittenmcinnische Zeituvg. 1863, p. 196 ; " Official 
Repot t:'- OesterreichiscJie Zeitschrift. 1889, p 297. 

2 Annales des Mines, 1845, viii., p. 293. 



SMELTING IN THE REVERBERATORY FURNACE. 



85 



is at c ; the opening /"carries off any lead fumes into the main flue. 
The hearth terminates at its lower end in a cast-iron gutter, over 
which the lead runs into the mould. Figure 22 shows the support for 
the working tools. The furnace is built of sandstone and ordinary red 
brick; the working bottom, which is renewed every four or five weeks, 
is made by tamping down firmly fire-clay, probably a mixture of 



Fig. 22 
FRONT ELEVATION. 




1 Vienna foot=-12.45 inches in Engliah 



raw and burnt clay. It is made impervious to lead by glazing 
With gray slag. The heating up of the furnace, leading to the frit- 
ting of the slag, is done slowly to prevent the cracking of the 
tamping. Furnaces are usually built in pairs, being placed side by 
side. They last from live to six years. The fuel used is cord- 
wood. Dimensions are given in the figures and in § 45. 



86 METALLURGY OF LEAD. 

The mode of operation is as follows : the furnace, barely red- 
hot from a previous charge, is repaired, if necessary, and the charge 
of about 400 pounds introduced through the working- door ^, and 
spread out over the upper part of the hearth near the bridge to a 
thickness of from l£ to 1^ inches. No fresh fire is made, the heat 
of the furnace and the glowing fuel from the previous charge 
furnishing sufficient heat for the first slow roasting. The ore, con- 
taining a small amount of blende, is raked every quarter of an 
hour. With pure galena the raking is not repeated so frequently, 
as the quick oxidation would liberate sufficient heat to make the 
ore cake. The beginning of this is recognized by its adhering 
to the rake. The roasting period has come to an end when 
the blue sulphur-flames disappear, drops of lead are seen near 
the front end of the hearth and the ore has a sandy feel. The 
roasting period lasts three hours, during which time the charge is 
rabbled from eight to nine times and from five to six sticks of cord- 
wood are consumed. The fire is then urged, to raise the tempera- 
ture. This is kept pretty uniform throughout the reaction, which 
now sets in. The attendant works his charge once every quarter 
or half hour, and raises the temperature slightly when the flow of 
lead ceases. This second period lasts from five to six hours, con- 
sumes from sixteen to eighteen sticks of wood, and furnishes the 
first half of the lead, which, on account of its freedom from im- 
purities, is called virgin lead. The attendant now stops firing 
until he has collected the residue from all parts of the hearth into 
one heap. He then takes a few shovelfuls of ashes and breeze 
from the ash-pit, throws it on the heap of residue, and works it in in 
order to remove lead and to reduce lead oxide and sulphate. He 
then urges the fire as quickly as he can, and the so called third 
period of the process, that of slag reduction, has begun. The 
further manipulations are the working of the residue and the stir- 
ring-in of breeze of charcoal, until after three hours the rest of the 
lead, the " slag-lead," has been extracted. This has to be liquated 
before it can be marketed. 

The practice is varied in some works by raking out the residue 
after the ashes and breeze have been stirred in and introducing a 
new charge. This is worked in the usual way. The residue from 
this second charge is not withdrawn, but that from the first added 
and both worked together for slag-lead. In this case the reduction 
of the slag occupies from seven to eight hours. 

The final residue is withdrawn from the furnace and sorted into 



SMELTING IN THE REVERBERATORY FURNACE. 87 

gray slag with 4 per cent, of lead, which is thrown away, and a 
product to be crushed and concentrated ; the heads which assay 
from 50 to 60 per cent, of lead, going back to the furnace in one of 
the subsequent charges of the residue. 

There is one furnace-man, working twenty-four hours, who has 
a helper during the day (12 hours). 

Tabulated results are given in § 45. Of the products no satis- 
factory analyses exist except of the lead. 





Virgin-Lead. 


Slag- Lead. 


a 1 


b 2 


a 1 


b 2 


Cu 


0.00069 
0.00025 
0.00055 

0.00076 

0.01476 
trace, 
trace. 


0.00010 
0.00003 
0.00700 

0.00400 
0.05700 


0.00075 
0.00025 
0.00088 
trace. 
0.00082 
0.01785 
0.00703 
0.00721 


0.00100 
0.00008 
0.00770 

0.00130 
0.14340 
0.01920 


A 2" 


fr& 

Fe 


Ni 


Zn 


S 


Sb 


As 





1 Otsterreichisches Jahrbudi, vol. xxii., p. 1189. 

2 Oesterreichisches Jahrbuch, vol. xxvii., p. 188. 



B. Lead- Smelting at JEngis, 1 Belgium. — The method of work- 
ing at Engis differs in some respects from that at Raibl. The 
ores are pure. They contain, according to an average of several 
analyses : PbS, 93.5G ; ZnS, 3.74 ; FeS 2 , 2.31 ; CaC0 3 , 0.35 ; traces 
of silver, and are free from arsenic and antimony. 

The furnace has the ordinary form of a reverberatory. The 
hearth, widest in the middle, is slightly contracted at the bridge and 
narrows down considerably at the flue. The furnace has two doors, 
oneat the side and one beneath the flue, below which there is a small 
kettle with a separate fireplace to receive the lead. The fuel used is 
bituminous coal. The furnace hot loin, oval in cross-section, begins 
at the top of the bridge, where it is •_' feet 8 J inches thick, and is 
inclined towards the flue, so that no Lead 18 collected in the furnace. 
At the flue the thickness is 1 foot 5J inches. This is the only 
furnace on record having a brasqne working-bottom. The brasque 
consists of two parts by volume of ordinary brick clay and one of 

1 Bouhy : Annates den Mines, 1870, xvii., ]>. 150; reprinl : La fabrica- 
tion du plomb, Dunod, Paris, 1870; also in Berg- und H&ttenm&nnische 

Zeitnng, 1870, p. 381 ; 1871, p. 52 (I\ 



SS METALLURGY OF LEAD. 

coke ground fine enough to ])ass a 4-mesh screen. Old bottoms, con- 
taining usually about 2 or 3 per cent, of lead, are ground up and 
mixed with new brasque. All the brasque needed for one furnace 
bottom (about 4,000 pounds of clay and 1,600 pounds of coke) is 
prepared by two men in 24 hours. In addition to the ordinary test 
for the correct amount of moisture (that the brasque, when squeezed 
in the hand, shall cohere into a lump, but not have sufficient moist- 
ure to adhere to the hand) another one is used, that of throwing 
with force a ball of brasque against the wall, to which it should ad- 
here. In tamping a layer of brasque 8 inches thick is first spread 
out evenly on the brick hearth and tamped down to 1 inch ; the sec- 
ond 8-inch layer is not rammed down as firmly, as it is reduced only 
to 2| inches ; the subsequent layers are made by using smaller 
amounts of brasque, as when spread out they are only about 
4 or 5 inches thick. The bottom is not impervious to lead, which 
filters through to a slight extent, collects on t p of the under- 
lying red brick, and also passes into the joints. A bottom lasts 
about six weeks. It wears off quickly during the first two weeks, 
but then resists pretty well abrasion by the tools and chemical 
action. 

In starting a furnace with a new bottom the warming lasts 
six hours, slight cracks that may form being closed with brasque. 
Then a small charge of 220 pounds of low-grade ore with from 
45 to 48 per cent, of lead is spread over the hearth and the 
temperature raised gradually for five and a quarter hours, the ore 
being raked from eight to ten times. The brasque becomes 
superficially soaked with lead and coated with a mixture of 
various more or less melted lead compounds. There result 
from this charge 4 pounds of lead, 5 pounds of rich slag-like 
material, and 130 pounds of matte-like material containing 48 per 
cent, of lead. The charges are increased in weight and richer 
ore is worked until, after the third day, the furnace can do eco- 
nomical work. 

The method of working is the same as at Raibl, with the excep- 
tion that all the lead is not extracted in the reverberatory. The 
residue, forming 12 per cent, of the original charge and assaying 
from 17 to 20 per cent, of lead, is melted with puddle-cinder in 
a small shaft-furnace. One furnace-man with a helper works a 
charge in twelve hours. Tabulated dimensions and results are 
given in § 45. 



SMELTING IN THE REVERB ER AT ORY FURNACE. 89 

The subjoined analyses show the composition of the resi- 
due : 

Si0 3 29.86 28.10 

PbO 25.50 32.80 

ZnO 25.33 20.80 

(FeAl) 2 3 15.03 14.70 

CaO 3.70 3.20 

S 0.38 0.10 



99.80 99.70 

C. Lead-Smelting in the Air-Furnace} — The roasting and reac- 
tion process, as carried out in the so-called air-furnace, is given here 
as the American improvement of the Raibl furnace. 

The galena ore from the Mississippi Valley is concentrated to a 
high-grade pure product ranging from 70 to 84 per cent, in lead. 

The furnace has a peculiar construction. The drawings (E'igures 
25 and 26) show an inclined hearth c, with e as charging and work- 
ing door, and y as discharging and cleaning door ; beneath is a small 
kettle </, into which the lead flows as it is set free in the furnace. 
The chimney is a sheet-iron pipe. Thirteen inches from the front — 
i.e., the lower or pot end of the furnace — is the fire-bridge b> with 
fireplace a at right angles to the axis of the furnace. The bottom 
of the furnace is a cast-iron plate with 6 inches of gray slag or 
residue melted upon it. Cord-wood is used as fuel. The cost of a 
furnace, including shed, is estimated by Broadhead at $550. 

The charge, from 1,400 to 1,600 pounds of galena of from pea 
size to hazelnut size, is introduced through door e and spread evenly 
over the hearth. It is roasted for from one to one and a half 
hours at a low temperature, the shorter time being sufficient 
if the galena contains some oxidized ore. During the roasting 

1 Williams : " Geological Survey of Missouri : Industrial Report," 
1877, pp. 8-101 ; Trans. A. I. M. E., v., p. 314.— Broadhead : "Geo!. Sur- 
vey of Missouri,*' 1874, p. 492. The furnace has been used in the Missis- 
sippi Valley, where it originated, to a considerable extent ; but even ten 
years ago it had given way largely to the ore-hearth, as this has about 
three times the capacity, although it saves a smaller amount of lead. The 
percentage of lead ore worked in 1880 by different furnaces in the Missis- 
sippi Valley is shown by the following table, taken from the Tenth Census 
Qf the United States, vol, x\\, p. 818. 

Air-Furnace 10.01 per cent. 

Flintshire Furnace 7.64 " 

Scotch Hearth 62.47 

Blast-Furnace 19.28 " 



90 



METALLURGY OF LEAD. 



the charge is constantly raked and moved from fire-bridge to fine 
and from the cooler part of the furnace to the hotter, in order to 
heat and roast it as uniformly as possible. When this is accom- 
plished, the heat is raised and lead begins to flow. Ashes and 
breeze are used as stiffening ingredients. The charge is rabbled at 
intervals. When no more lead appears (after from seven to eight 
hours), the residue is drawn without any attempt to extract slag- 
lead in the furnace as at Raibl. 

One furnace-man with a helper works in twelve hours a charge 
which consumes one and one-half cords of wood. About the yield 
nothing is known, as the contents of ore charged is not taken. 
Tabulated dimensions and results are given in § 45. 

The following two analyses by Williams 1 show the composition 
of residue and poled lead from the Granby works : 



Fig. 25 

V 




VERTICAL SECTION ON LINE A B. 



Residue: Si0 2 , 21.396; CaO, 4.650; MgO, 3.948; Fe g O„ 
3.680; A1 2 3 , 0.152; ZnO, 7.146; PbSO* 2.349; PbS, 20.929; 
PbO, 34.914. Total, 99.063. The sample yielded 3.52 per cent, of 
metallic lead, which makes the total metallic lead 54.82 per cent. 

Lead: As, 0.01122; Sb, 0.00077; Ag, 0.00080; Cu, 0.05091; 
Fe, 0.01582 ; Zn, 0.00090 ; Ni, 0.00281 ; Pb (by difference), 
99.91777. Total, 100.00000. 

§ 43. The English Method. — The characteristics of this 
method are a large charge, a quick roast (with the result that for 
every part of lead sulphate formed there shall remain two parts of 
lead sulphide unchanged), a high temperature throughout, and the 
aim to extract all the lead in the reverberatory. The hearth in- 
clines towards the middle of one of the sides ; the lead collects in 
the furnace, and is tapped at intervals into an outside kettle. 



1 Trans. A.I. M. E., v., pp. 320-324. 



SMELTING IN THE REVERBEBATORY FURNACE. 



91 



Lead- Smelting at Stiperstones 1 (Shropshire). — The ore is a 
concentrated galena with some carbonate and blende, assaying 
77.5 per cent, of lead, the little gangue being principally carbonate 
of lime. 

The construction of the furnace i3 given in detail in Figures 
27 to 32. The horizontal section (Figure 27) shows the usual 
trapezoidal form of the English reverberatory with its three work- 
ing doors Jc on either side, the well b at the front, and the fire- 
opening t at the back. It is to be noted that centres of hearth a 
and fireplace g are not opposite each other, the latter, with fire- 
bridge A, being 10 inches further back. The object of this is to 




draw away the flame from the well b, for the same reason the 
flue d, near the back, is made 1 inch wider than flue d, while both 
have the same height. Each has its own damper w and w'. The 
cross-section, Figure 29, shows the inclination of the hearth from 
the back to the front (24 degrees), where well b is placed. This 
has nearly vertical walls and (Figure 27) extends 3£ feet into the 
furnace, where it is 2 feet wide, narrowing down to a few inches 
in the front. The lead collected in the well is tapped through tap- 
ping-hole m into the cast-iron kettle or pot c. The longitudinal 



1 Moissenet : Annalcn des mincx, 1862, vol. i., p, 15; reprint, "Traite- 
ment de la galene au four gallois," Dunod, Paris, lHf>2 ; also in Bery-uti<l 
Hiittenmamiische Zeitung, 1863, pp, 243, 251, 261, 166. 



92 



METALLURGY OF LEAD. 



section (Figure 28) shows the inclination of the hearth, which is 30 
degrees from the bridge and 23 degrees from the flue. The roof 
P slopes in a straight line from top of grate g to flues d d'. The- 
gases from the grate pass over fire-bridge h to the hearth and 
thence through flues dd' into flue e. In the front elevation 
(Figure 30) can be seen the chimney f, which is 2 feet 6 inches by 
2 feet in the clear and 20 feet high from the ground. It does not 
communicate with the interior of the fireplace. It carries off 
gases that would enter the smelter building when the fire is fed or 
stoked or when the doors are open, and also vapor that comes from 
the ash-pit a (Figure 28), which is filled with water. 



HORIZONTAL SECTION ON LINE R, S 




The following details will elucidate the drawings. In Figures 
27, 28 and 30 one end of the furnace is seen to be built against the 
wall, Q, of the building. At the others there is a passage, D; 1 5 
feet of free space at front and back give sufficient room for working 
the furnace. 

The furnace is built of red brick and lined with a full course of 
fire-brick, which is indicated by heavy shading. The foundation 
is 3 feet deep ; the space in the centre between the bent ends (j f 
in Figure 29) of the lower tie-rods is filled with sand. The upper 
part of the foundation is solid and slopes from the back and both 
ends to the front. The edges of the last course of brick are left 



SMELTING IN THE REVERBERATORY FURNACE. 



93 



sharp, as shown by the zigzag line in Figures 28 and 29. In 
determining the slope, a line 3 feet 6 inches long is drawn from the 
tap-hole, and the points a and P located at a distance of one foot 
on each side of its inner extremity. To them, lines are then drawn 
from the corners 7 and <*, starting at an elevation of 2 feet 8 inches. 
The side-walls are raised slightly over the roof P, and have a total 
height of 5 feet 8 inches, measured from the floor. The space be- 
tween the two is filled with sand, R'. 

The binding of the furnace is clearly shown in the figures. 
The buckstays are 6 feet 6 inches long, the tie-rods being slipped 
over them and tightened with wedges. The upper tie-rods reach 
over the furnace ; the lower ones, jj', are turned down one foot at 



Fig. 28 
LONGITUDINAL SECTION ON LINE X, Y, Z, V. 




7 






a distance of 4 feet 6 inches. The castings are not shown with 
sufficient distinctness in the drawings. 

The tap-hole plate (Figure 29 and y in Figure 30) is 5 feet long 
and 20 inches high; it is two inches thick at the bottom, increasing 
to three, which it reaches at a height of 6 inches and retains for 
the rest of the distance. Eight inches above the centre of the 
lower edge is the tap-hole ???, and 4 inches above this begins a 
narrow vertical opening (4 by 9 inches), with a hinged door, p, 
only R inches high, so that when the door is closed a small current 
of cold air may pass over the molten lead in the well. The tap- 
hole plate is overlapped or each side by live east -iron plates (4 feet 
long and -£ inch thick), placed horizontally (z and z' in Figure 



94 



METALLURGY OF LEAD. 



30). They extend from the working opening down to the tap- 
hole. 

The back-plate (5 feet long, 20 inches high and 1 inch thick) is 
placed just opposite the tap-hole plate. Its main object is to pro- 
tect the brick from the gray slag, which is drawn out from the 
back door. To facilitate this, two hooks, II (Figure 29), carrying 
an iron rod (3 feet 6 inches long and 1 inch in diameter) as sup- 
port for the working-tools are placed inside of neighboring 
buckstays. 

The bridge-plate n (Figure 28), counteracting the longitudinal 
thrust of the hearth, gives strength to the 4-inch air-space o, and 
prevents any leakage of lead. It is 6 feet long and 20 inches high. 

Fig. 29 
CROSS SECTION ON LINE T, U. 




From the upper edge, which is on a level with the grate bars, it is 
3 inches thick for a distance of 6 inches, then suddenly increases 
to double the thickness for another 6 inches, and for the last 8 
inches returns to its original thickness. 

The working-openings, k, have a number of castings. The 
door-frames (B, Figure 29), enclosing each an opening 10 by 6 
inches, are 4 inches square in cross-section. They are protected 
on the sides by fire-brick, and on the bottom by the hearth. On 
the level of the grate-bars, or 2 feet 8 inches above the floor, are 
two horizontal plates, C, 6 inches wide and \ inch thick, that ex- 
tend below the six working-openings. On them rest two inclined 
(3 : 8) plates, E, 8 inches wide and \ inch thick, which abut against 



SMELTING IN THE REVERBERATORY FURNACE. 



95 



the bottom of the door-frames. Two plates, F, of the same size as 
C, form the upper part of the linings. The skew-backs, G (9 inches 
wide and | inch thick), resting on these upper plates, F, and the 
door-frames, B, support the roof. The sides of the working-open- 
ings are lined with £-inch jamb-plates, as shown in Figure 27. The 
cast-iron working-doors, (/, with handle, r (Figure 31), are placed 
against the door-frames, B. 

The cast-iron kettle, c, in front of the tap-hole is 2 feet 6 inches 
in diameter and 20 inches deep. It is embedded in clay, which is 




^\\\K^0 



enclosed on three sides by masonry and on the fourth by the slag- 
bottom. Its rim, 8 inches above the floor, is surrounded by an iron 
hoop, 4 inches wide and £ an inch thick. Between it and' the kettle 
wedges, T, are driven to prevent any lead that may penetrate 
through the joints and collect under the kettle from raising it up. 
The fire-opening, t, 10 by 12 inelies, has an iron casing (Figure 
27). It is closed by a swinging fire-door (Figure 32), consisting of 
fire-brick held together by a wrought-iron frame. The grate 
consists of two sets of cast-iron grate-bars of eight each, sup- 
ported at the ends by cast-iron cross-pieces. 



96 METALLURGY OF LEAD. 

The hopper (x, Figures 28 and 30) is a truncated sheet-iron 
pyramid 36 and 6 inches square at the two ends and 3 feet high. 
It is suspended in a wooden frame, K (3 inches square), which rests 
on two iron cross-beams, L, 4 by 6 inches. These are supported 
by brickwork, 14 inches high, forming the continuation of the side- 
walls of the furnace. Four iron rods, N (Figure 28), fastened in 
the wooden frame, K, have at their lower ends a sheet-iron frame, 
O, on which is placed a movable slide, which closes the discharge 
of the hopper. The 6-inch opening in the roof of the furnace, over 
which the mouth of the hopper is placed, is indicated on the hearth 
in Figure 27. It is 1 foot 9 inches distant from the fire-bridge, 
and 2 feet 6 inches from the working-opening nearest the bridge 
at the back of the furnace. 

When the furnace is finished, the working-bottom, made of 
coarsely-broken gray slag and sand, is put in. Part of the mate- 



WORKING DOOR. 




^^ 0fc=] 



SCALE FOR FIGS. 27. 28, 29 AND 30. 

f i i i f £ F* \" 

, SCALE FOR FIGS. 31 AND 32. , , 

flMlllllllll 1 - f 1 t 

rial is spread out on the brick bottom, which has previously been 
made red-hot. It is heated till it becomes pasty, and is then patted 
with paddle and rake. A second part is made to adhere to the first, 
and so on until the entire bottom has been built up with successive 
layers of gray slag and sand, and has attained the desired form 
and thickness. The upper edge of the well is 10 inches below the 
working-opening. The thickness of the working-bottom increases 
towards the tap-hole. It is 4 inches at the back doors, 12 at the 
bridge, 14 at the flues, 30 at the front, and 8 at the back of the 
well. 

The furnace, being built solidly, lasts a long time ; the roof of 
the fireplace requires repairing every two years, that of the hearth 
every five years. The hearth is repaired after every charge if 
necessary. 

This furnance has the representative form of the English rever- 



SMELTING IN THE REVERBERATORY FURNACE- 



97 



beratory. Slight variations exist as regards detail. For instance*, 
the foundation, instead of being built up solid at the back, some- 
times has an arched vault x at J (Figure 3), extending longitudinally,, 
communicating with the air-space in the fire-bridge, and being acces- 
sible at the flue end of the furnace. Another variation is when the 
entire hearth is built on cast-iron plates 2 resting on rails sup- 
ported by a brick pillar. Sometimes the general form of the 
hearth differs by having a gentle slope towards the tap-hole. This. 
makes the sides of the well less steep than those of the drawings. 

The tools used in the furnace are shown in Figures 33 to 46- 
They are wrought-iron, with the exception of the mould (Figure 45) 
and the handles of shovel (Figure 37) and ash-pit hoe (Figure 40). 

Figures 33 to 36 : Paddle, rake, old paddle used as chisel, ham- 



FIREl 



DOOR 



-Fig. 32 



ijF 



mer to remove adhering slag, are the tools for working on the 
hearth. Figures 'S7 to 40 : Coal-shovel, hammer to break coal, 
fire-poker, ash-pit hoe, are required for firing. Figures 41 to 46 : 
Tapping-bar, rectangular skimmer, circular skimmer, ladle, mould, 
lead-carier, are used in handling the lead. 

Method of Working. To simplify the description the doors 
may be be designated by numbers 1, 2, 8, in front, and 4, 5, 6, at 
the back, Starting both times from the bridge. 

The charge, 2,350 pounds, is let fall from the hopper through 

the roof into the furnace, still red-hot from a previous charge, the 
dampers having been closed. The lead is left in the well and 
covered with lime. The ore is spread over the hearth through 

1 Perry, "Lead," pp. 282-889. 

2 Phillips-Baurrman, Op, d$. f p, 591. 



98 



METALLURGY OF LEAD. 



doors 1, 4 and 5, with rakes. It decrepitates and gives off vapor 
of water. Then the dampers are slightly raised and the fire is 
gradually increased for an hour and a half. During the first two 
hours (first firing) the ore is turned over four times with padddles 
through doors 1 and 4, the other doors being kept closed. The 
paddling requires only a few minutes. When the first firing is 
nearly accomplished, feeding of fuel is stopped, and doors 1, 2, 4 
and 5, fire-door and dampers are thrown open and the charge 
cooled (first cooling). This lasts for half an hour. The grate is 
then freed from clinkers, and the charge, nearly 6 inches deep next 
to the bridge, which has fritted, is broken up and turned over. Any 
ore that had fallen into the well is raked up on the furnace-bed. 
The doors are now closed, the dampers lowered, the grate is well 
filled with coal, and the second firing begins. This lasts from 55 



WORKING TOOLS. 

Fig. 38 



Fig. 37 



*q: 



Fig. 33 




B Fig. 41-==== 



,FlG. 

39 




to 60 minutes, and is followed by a second cooling. Some lead 
now flows into the well. Near the bridge and towards the centre 
of the furnace, parts of the charge have begun to fuse, while at the 
■flue the ore is only sintered. The charge is worked as in the first 
cooling, door 6 now being also used. After ten minutes the lead 
which has accumulated is tapped from the well into the kettle. 
The tap-hole is closed from the inside by inserting a clay plug fixed 
to a wooden handle through the little tap-hole door and pressing 
it in until the clay oozes out in front. Before tapping, fine coal 
and wood-shavings are put into the kettle to pole the lead. It is 
then stirred vigorously for six or seven minutes with the rectangu- 
lar skimmer to bring all the impurities to the surface. These are 
removed, first with the shovel and then with the rectangular skim- 
mer, and thrown back into the furnace, through doors 1 and 3, on 
both sides of the well. The charge, now freed from part of its 



SMELTING IN THE REVERB ER AT ORY FURNACE. 99 

lead, is turned over with paddles through doors 1 and 4. This 
ends the second cooling, four hours after charging. The doors are 
now closed, coal is put on the grate, and the dampers are lowered 
to begin the third firing. Meanwhile the lead in the outside kettle 
is skimmed and ladled into moulds holding one hundred and twenty 
pounds. This takes about twenty minutes. During the third 
firing, which lasts two hours, the ore is turned over several times, 
care being taken to open the doors as little as possible. The fur- 
nace shows a bright-red heat when the third cooling begins. The 
charge is now worked with paddles for fifteen or twenty minutes, 
and parts of it that have collected in the well are stiffened by the 
addition of lime and are raked on the hearth. The residue on the 
hearth is collected near the bridge and fine coal worked into it. 
The doors are again closed, the fire is urged for a quarter of an hour, 
and the residue turned over with the paddle through doors 1 and 
4, and finally drawn out through door 5. Any repairing of the 
hearth that may be necessary takes place now, and the furnace is 
ready again for a new charge, seven hours from the time when the 
previous one was first introduced. 

Two men work as partners in twelve-hour shifts. Tabulated 
results are given in § 45. 

The method of working near Holywell, North Wales, differs 
from that at Stiperstones. According to Percy * it is as follows. 

After dropping the charge of 2,350 pounds on the hearth and 
spreading it over the upper part of the bed, the doors of the fur- 
nace and that of the fire-box are left open for an hour and a half 
to let the air have free access during the first roasting stage, while 
the damper is raised just enough for the gases to escape. Work- 
ing-doors 3 and G and the fire-door are now closed and the fire is 
urged. Lead soon appears. During this heating, which lasts two 
hours, doors 2 and 5 are closed, but 1 and 4 are kept open. 
Through these the charge is rabbled at intervals. Towards the 
end of this first reaction period lead begins to flow. Now doors 
1 and 4 are closed, the damper is thrown open, the tire urged for 
forty minutes, and the charge melted down. The furnace is then 
cooled for half an hour by throwing open all the doors. What 
charge remains on the hearth is rabbled, d<>nrs :; and •'» are closed, 
slacked lime is thrown through door i <>n the charge, which has col- 
lected in the well, and worked into it through the tap-hole door. 

1 "Lead," p. 232. 



100 



METALLURGY OF LEAD.. 



The stiffened residue is collected (" set up ") near the bridge as well 
as other parts that have been detached from the hearth. Doors 2 
and 5 and the lire-door are now closed, the damper is lowered, and 
the temperature raised gradually for half an hour, when the 
damper is entirely thrown open, doors 1 and 4 are closed, and the 
fire is urged to melt down the residue, which takes twenty min- 
utes. The fire-door and the working-doors 1, 2, 4 and 5 are 
then thrown open, lime is added through door 2, and worked into 
the slag to thicken it. The lead is tapped, and the stiff gray slag 
raked out on the floor through door 5. The lead is poled as at 
Stiperstones, and the hearth repaired, if necessary. The entire 
time required to work the charge is six hours. 

Two furnace-men and one helper work two charges in a twelve- 
hour shift. Tabulated results are given in § 45. 

The changes that take place in the ore during the process have 
been examined by Percy, 1 and are given here in their lead contents 
only. 



Lead Com- 
pounds. 


Hours after Charging. 


Gray Slag. 


1^ 


3^ 


VA 


m 


PbS 

PbO 

PbS0 4 


63.82 
27.25 

3.82 


53.32 
31.49 

4.78 


24.76 

43.12 

6.94 


4.35 
47.50 
14.02 


0.90 

48.87 
9.85 


Total.Pb, 


83.16 


78.66 


66.22 


47.86 


52.88 



§44. The Silesian Method. — The characteristics of this 
method are a large charge, slow roasting, and a low temperature. 
It is not aimed to extract all the lead in the reverberatory, as this 
is supplemented by the blast-furnace. The hearth is inclined 
toward the flue, beneath which the lead is collected and tapped 
at intervals into an outside kettle. 

Lead-Smelting at Tamowitz? Prussia. — The ore is a mixture 
of sulphide and oxide lead minerals. Its composition was in 1870, 
in round numbers : 



1 Op. cit., p. 235. 

2 Zeitsehrift fur Berg-, Hiitten-, und Salinen-Wesen in Preussen, vol. 
adv., p. 225, Teichmann ; xix., p. 157, Wedding ; xxxii., p. 94, Dobers and 
Dziegiecki ; xxxiv., p. 292, Dobers and Althans. 



SMELTING IN THE REVERBERA TOR Y FURNA CE. 101 



PbS 61 

PbS0 4 11 

PbC0 3 24 

(Ca, Mg, Fe, Zii) C0 3 3 

SX>„ .... 1 



40 

9 
45 

5 

1 and less. 



The furnace-charges contained from 70 to 74 per cent, of lead 
and from 21 to 22 ounces silver per ton. At present the ore runs 
lower, the percentage of blende having increased considerably. 

The construction of the latest furnaces is given in Figures 47 
to 50. After the detailed description of the reverberatory at Stip- 
erstones attention need only be called to some of the principal 




Id ^°- 

HORIZONTAL SECTION ON LINE G, H. 



features. The horizontal section (Figure 47) shows the rectangular 
form of the furnace with four working openings on either side, 
the well being below the door nearest the flue end, the coolest 
part of the furnace. The grate bars are placed parallel with the 
axis of the furnace. The fire is stoked at the end and fed from 
both sides. Four branch flues lead the gases into one main flue. 
The bridge-wall has an air-passage. The entire furnace is encased 
in iron plates and bound by backstays (iron rails) and tie-rods. In 
the longitudinal section, Figure 48, the roof is seen to form a 
horizontal line from the fireplace to the beginning of the hearth, 
.loping thence in a straight line to the tines. The hearth, in Fig- 
ures 48 to 50, is built up differently from that of any other furnace. 



102 



METALLURGY OF LEAD. 



First a layer of sand is tamped down between the furnace-walls so 
as to have the form of a trough inclining from bridge to flue. At 
the well the brick bottom, of red-brick set dry, replaces some of 
the sand. It forms a good support for the brasque bottom, which 
is followed by the slag bottom, consisting of tap-cinder melted 
down in the furnace. This covers the brasque bottom, wherever 
it would otherwise come in contact with working-tools, leaving it 
exposed only at the well (Figure 49). During the run some residue 
adheres to this slag bottom and forms the smooth working bottom 
of the furnace. The hearth is seen in Figure 50 to be trough- 
shaped from the bridge to the second working-door ; from there 
the front part (Figure 49) slopes down to the level of the rim of 
the kettle, the lowest part of the well. 

The tools required to work a furnace are four paddles, four 
large and two small rabbles, five shovels (two for lime, two for 



TZ7H Fire brick 
Red brick 



LONGITUDINAL SECTION ON LINE A, B. 




coal, and one for slag), three steel bars (two large ones and a small 
one), a tapping-bar, a skimmer, a ladle, a sample-ladle, two slice- 
bars, a sledge, tw T o hammers, and four hooks for handling the lead. 
Method of Working. — The furnace, if new, is heated to a dark- 
red heat ; the damper is then closed and the charge, crushed to 
pass a 5-niesh sieve, is let fall from the hopper through the opening 
in the roof and spread out evenly over the hearth by means of 
rabbles to a thickness of from 3 to 4 inches. The fire is fed with 
cinders and the temperature never allowed to exceed dark-redness, 
say 500 or 600 degrees Centigrade. The galena decrepitates, the 
temperature rises, the ore becomes a dark-red, and after from 
three-quarters of an hour to an hour the roasting begins. During 
this time the charge is turned over once with the paddle. The 
w r orking-doors and fire-door are opened, and the damper raised 
sufficiently to allow the sulphurous acid and other gases to pass 



SMELTING IN THE REVERBERATORY FURNACE. 



103 



off. The ore roasts on the surface. When the fumes begin to 
cease, samples are taken to see if a white crust of oxide and sul- 
phate has been formed. This indicates that it is time to renew the 
surface, which is generally done by paddling or rabbling every 
twenty or twenty-five minutes, i.e., about eight or nine times dur- 
ing the three or four hours required for roasting. Care is taken to 
prevent the ore from clotting. 

Up to 1885 the roasting usually took about three or three 
and a half hours, but the gradual increase of blende in the ore 
brought the required time up to four hours. In 1886 it became 
necessary to make a change in working the charge. This could 
be done in two ways. Either the normal time could be prolonged 
or part of the ore roasted separately. The latter method was 
chosen, and now one-quarter of the charge, fine concentrates es- 
pecially rich in blende, is roasted separately in a long-hearth cal- 



VERTICAL SECTION ON LINE E, F. 




SSS^J^i--. ^c> 



cining furnace and then added to the ore charge. Towards the 
end of the roasting period from 330 to 660 pounds of carbonate ore 
or fluedust, with 45 per cent, of lead, are added to increase the 
amount of oxides in the charge. 

The grate is now cleaned, well filled with good coal, and the 
damper opened to raise the temperature. The reactions now se1 
in. The roasted ore gradually softens, white fumes are given oil", 
and lead begins to flow. Lime is added to prevent the lead from 
carrying down to the well particles of ore floating on top of it 
and to prevent liquefying of the charge. The lime is well 
worked in and the charge then turned over. The furnace be- 
comes filled with fumes of lead, the damper is raised to remove 
these, ami fusing of the charge is prevented by regulating the 

fire and the damper and by adding a few shovelfuls of lime al a 
time. At from an hour to an hour and a quarter after the re- 
action has set in (three hours with a new furnace) the well is 



104 



METALLURGY OF LEAD. 



full of lead. This is tapped into the outside kettle. The dross 
floating on the surface is skimmed off and put back into the fur- 
nace and the impurities held in suspension by the lead are removed 
by stirring in slack coal (poling). This second dross is put aside 
and the lead ladled into moulds. During the reaction stage the 
working-doors have to be open on account of the necessary rab- 
bling and paddling ; thus air enters, cools the furnace, oxidizes the 
charge and the flow of lead begins to cease. 

At this stage drosses obtained from melting down base bullion 
in the desilverizing plant are added. They consist principally of a 
mixture of lead sulphide and lead, assaying 75 per cent, lead and 9 
ounces silver per ton. At the same time 1,100 pounds of oxides 
low in silver, containing from 75 to 80 per cent, lead oxide and 
from 19 to 20 per cent, zinc oxide, are charged. They result from 

ITig.SO 

VERTICAL SECTION ON LINE C, D. 



G * 




refining lead, desilverized by the Parkes process, by means of 
steam (§ 107). 

The doors are then closed and fuel is piled up high on the 
grate to insure a smoky flame. This reduces lead sulphate to sul- 
phide, and soon the flow of lead begins again. The doors are 
opened, the charge is worked, and its liquefying prevented as be- 
fore. From two and a half to three and a quarter hours after the 
first reaction sets in the well is again filled with lead. This is 
tapped and lead and dross are treated as before. In this way all to- 
gether five reactions are caused, the amount of lead obtained be- 
coming gradually smaller and the charge more dry. Before the 
last reaction takes places, the mixture of dross and fine coal, ob- 
tained from poling the different leads in the kettle, is thrown into 
the furnace. It is worked into the charge, which is then covered 
with fine coal. The temperature is raised and the last yield of lead 
obtained. The damper is now opened fully and the residue drawn 
from the back of the furnace into a box filled with water, so as to 



SMELTING IN THE REVERBERATORY FURNACE. 105 

prevent fumes of lead from passing into the building. Care has to 
be taken to avoid explosions. The hearth is repaired with a mixture 
of crushed residue and slacked lime, which is beaten down with the 
paddle and rammed into holes with a bent bar. When this is fin- 
ished, a little lime is spread over the hearth and a new charge drop- 
ped from the hopper. The whole reaction period lasts seven hours. 

If the hearth should become so incrusted as to raise the charge 
too high to be protected by the fire-bridge from the direct action 
of the flame, the temperature must be raised so as to soften the ac- 
cumulated residue that it may be removed. A hearth lasts about 
three months. 

The results, from one year's (1865) work in a smaller furnace, 11 
feet 9 inches by 10 feet 10 inches, with six doors, and with ore low 
in blende and without the different additions, were : 10,000 pounds 
of ore (assaying 72.97 per cent, lead and 21.76 ounces silver) gave 
6,384 pounds base bullion, assaying 32.95 ounces silver ; 1,592 
pounds residue (38.80 per cent, lead and 3.93 ounces silver) ; 275 
pounds fluedust (50.00 per cent, lead and 0.02 ounces silver) ; 
showing that in the reverberatory 87.49 per cent, of all the lead 
and 99.9156 per cent, of all the silver was recovered in the form of 
base bullion. Of the 12.51 per cent, of lead, forming the differ- 
ence, 9.31 were recovered in the blast-furnace, so that the actual 
total loss amounted only to 3.2 per cent. 

The consumption of fuel was 4,600 pounds, or 46 per cent, of 
the ore charged, and of lime 100 pounds or 1 per cent. Additional 
data are given in the next paragraph. 

The following analysis of residue was made by Pietsch in 1865: 
PbO, 24.375; PbS0 4 , 13.269; PbSi0 3 , 12.373; ZnO, 22.857; 
FeO, 8.957 ; FeS, 1.823 ; CaO, 11.190 ; C, 4.821 ; Al 2 O s and MnO, 
traces ; silver, 0.015 (= 4.360 ounces per ton). Total, 99.680. 

The fluedust, in 1881-1882, forming 2.91 per cent, of the ore 
charged, had the following composition, according to Dobers and 
Dziegiecki : 

PbO G6.44 66.53 86.79 

ZnO 3.77 4.55 4.05 

Fe 8 3 0.50 o.3() 0.85 

SO, 27.48 27.11 27.7!) 

Insol 2.14 1.20 3.80 



100.:;:! 99.69 100.28 

§ 45. Comparison of Reverberatory Methods.— T<> facili- 
tate comparison, the main data of the furnaces discussed have been 
brought together in the subjoined table: 



IOC 



METALLURGY OF LEAD. 



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108 METALLURGY OF LEAD. 

In comparing the amounts of ore treated in twelve hours at the 
different smelting works, the order in which they are placed in the 
table shows a steady increase from Raibl to Tarnowitz. The fig- 
ure for Tarnowitz, 8,250 pounds, requires the explanation that the 
charge contains a considerable proportion of oxidized ore, which 
shortens the time required for roasting. If the ore were pure ga- 
lena, twice the time, or eight hours, would have to be allowed. This 
would make the amount for twelve hours 6,187 pounds. The large 
amount of ore treated is due to the size of the furnace. As to the 
amount of labor required per ton of ore, the table shows that 
Tarnowitz uses less and Raibl more than any of the other smelting 
works. With fuel the same is the case. As the wear and tear of a 
furnace depend on the height of the temperature, the Carinthian 
and Silesian furnaces will outlast the English ones, other things be- 
ing equal. If, finally, the amount of lead recovered is considered, it 
is clear that slow roasting yields more than quick roasting (Ca- 
rinthian and Silesian vs. English), and that recovering in the rever- 
beratory only that amount of lead obtainable at a low temperature 
and not melting the charge, gives a higher percentage of lead (Tar- 
nowitz and Engis vs. Raibl and English methods). The inferences 
to be drawn are clear. They are stated by Cahen, 1 and Gruner 2 has 
formulated them at length. He says (Percy's translation 3 ) : " Rich, 
pure, non-quartzose ores ought always to be treated by this latter 
(the roasting and reaction) method. The operation ought to take 
place in large reverberatory furnaces, with easy access of air, pro- 
vided with a single fireplace and a receiving basin, internal or exter- 
nal, placed in the least heated region of the furnace. The operation 
ought always to be conducted slowly, and to consist of two phases 
very distinct, roasting and reaction. For roasting, the layer of ore 
must never exceed from 3.15" to 3.54" in thickness. Roasting 
is to be effected at a low temperature, and ought to proceed as far 
as the theoretical limit of one equivalent of sulphate, or two equiv- 
alents of oxide, for each equivalent of sulphide. After the first 
firing, which produces lead, and fresh roastings and firings twice 
or thrice repeated, the rich residues must be withdrawn from the 
reverberatory furnace, without having recourse to ressuage (re- 
duction of rich residue in the same furnace immediately after- 
wards), but rather by practising this ressuage in a blast-furnace." 

1 Op. cit., p. 117. 

2 Op. cit., reprint, p. 511. 

3 "Lead," p. 491. 



SMELTING IN THE REVERBERATORY FURNACE. 109 

Why in our own lead districts the reverberatory practice has 
not made more headway is to be answered in two ways. In the 
lead-silver districts the ores have on the whole not been of suffi- 
cient purity and richness to warrant the use of reverberatory 
methods. In the Mississippi valley, where the ore is just of the 
quality required for the process, the question of skilled labor has 
had some influence, but whether justly so is veiy doubtful. 












CHAPTEE VII. 
SMELTING IN THE ORE-HEARTH. 

§ 46. Introductory Bemarks. — The process carried on in the 
ore-hearth is the roasting and reaction process. It resembles that 
in the reverberatory furnace, with this difference, that oxidation 
and i*eduction go on simultaneously, the charge floating on a bath 
of lead. The lead oxide and sulphate, as soon as formed, react on 
undecomposed sulphide, and the liberated lead trickles through 
the charge into the hearth-bottom, overflowing into an outside 
kettle. 

The same conditions are necessary for the hearth-treatment as 
for the reverberatory, with the exception that the ore should be 
coarser. The smallest permissible size is that of a pea, and nut- 
size is desirable. If fine ore is to be treated, it must first be 
agglomerated in a reverberatory, as in the ore-hearth it would be 
blown away. It requires power and a blower. Then much lead is 
volatilized ; hence it is not suited for argentiferous ores. The com- 
parative loss at Raibl 1 where ore-hearth and reverberatory work 
on the same ore, was, in 1888, 10.6 to 7.3 per cent. It requires purer 
and higher grade ore than the reverberatory, but consumes less 
fuel. According to Tunner, 2 a furnace similar to the one at Raibl 
consumed, per 100 pounds of galena, 7.31 cubic feet of wood, while 
the ore-hearth required only 2.90 cubic feet. The ore-hearth has 
three times the capacity of the air-furnace. The cost of treatment 
per ton of ore is the same with a single ore-hearth ; with several 
running side by side it becomes less, as only one set of men is 
necessary to run the more powerful engine, and this consumes rel- 
atively less steam than the smaller one. That the ore-hearth can- 
not compete with the English or Silesian furnaces as regards capaci- 
ty and cost, is clear. It has, however, one advantage over all 
reverberatories, that it is quickly started and stopped without 

1 Oesterreichische Zeitschrift fur Berg- u. Hutten-Wesen, 1888, p. 329. 
8 Leobener Jahrbuch, 1852, i., p. 262. 



SMELTING IN THE ORE-HEARTH. 



Ill 



much consumption of fuel or loss in heat, and thus serves its purpose 
in extracting at intervals from small amounts of non-argentiferous 
ore the major part of the lead. This is probably the reason why it 
found such favor in the Mississippi Valley, where small amounts 
of ore were often treated, and still are, by men who have mined it 
themselves. 



:Fi g .51 

VERTICAL SECTION ON LINE G, D. 




\d 



FRONT ELEVATION. 




Fig. 53 
HORIZONTAL SECTION ON LINE A, B. 




The ore-hearth as worked at Joplin, Mo., is exceptional and 
will be treated further on ($ 51). 

The practice in the different ore-hearths is so very much the 
same that nothing from a general point of view need be said 
about it. 

The products are similar to those of the reverberator}' ; there 



112 METALLURGY OF LEAD. 

is, however, an intermediary product, a mixture of ore, slag, and 
fuel, called browse in England, which goes back to the charge in 
the ore-hearth. The slag contains much lead, and is smelted in a 
small blast-furnace, the slag-eye furnace. 

§ 47. Influence of Foreign Matter.— It has already been 
said that the ore subjected to hearth treatment must be purer 
and richer than is necessary for the reverberatory furnace. This is 
because the foreign matter shows its bad influence in a more 
marked degree. The chemical action is, however, the same as that 
described in § 40. 

§ 48. Description Of Ore-Hearths.— An ore-hearth, being 
a small low fireplace surrounded by three walls, with a tuyere at 
the back, cannot show much variety in construction or in manner 
of working. Three slightly different forms have been chosen by 
way of illustration : 

The Scotch Ore-Hearth, 

The American Water-Back Ore-Hearth ; and 

The Moffet Ore-Hearth. 

a. The Scotch Ore- Hearth. 1 — Figures 2 51-53 represent the fur- 
nace used by Messrs. Cookson and Co., near Newcastle, in the 
North of England. The cast-iron hearth-box or well a, which 
holds the lead, is set in brick-work q. It is 2 feet from front to 
back and about 2 feet 6 inches wide ; it is 1 foot deep and holds 
about two tons of lead. In some furnaces the depth is only 6 inches 
and the capacity of the well about 1,340 pounds of lead. The 
work-stone (/, an inclined plate, is cast in one piece with the hearth- 
box. It has a raised border on either side and at the lower edge, 
and a groove h, which leads the overflowing lead towards the ket- 
tle i, heated from a fireplace J, below, the gases passing off through 
a flue into the chimney /. On either side of the hearth-box and 
resting on it is a cast-iron block (bearer) n. Another cast-iron 
block o (back-stone, pipe-stone) is placed at the back. It is per- 
forated for the passage of tuyere b, which enters the furnace about 
2 inches above the surface of the lead in the well. Upon o rests the 
upper back-stone />, also of cast-iron. Thus lead and ore are 
entirely surrounded by cast-iron. The fore-stone m appears to be 
rather small. The brick shaft c carries off the fumes ; at the 

1 Percy, " Lead," p. 278. 

2 '* Eighth Annual Report of the Local Government Board," 1878-79, 
supplement containing the Report of the Medical Officer for 1878, London, 
1879, p. 281. 



SMELTING IN THE ORE-HEARTH. 



U1& 



back is a "blind flue or pit" d, forming a sort of dust-charnbeiy to* 
be cleaned through door e, the gases passing off upward. On the- 
side of the shaft is the feed-door k, for the introduction of fuel im 
front of the tuyere and the removal of any slag adhering to it, the- 
charge being fed from the front. The hearth has at the front a 
shutter f, sliding in a grooved frame r / it is raised and lowered 
by means of counterpoise s. Peat, formerly used as fuel, is now 
replaced by bituminous coal. 

It is characteristic for the method of Avorking in the Scotch 
hearth that the process is non-continuous. After smelting from 
twelve to fifteen hours the hearth becomes too hot ; it has to cool for 
about five hours before work can be started again. 

VEETICAL SECTION. 

<= 7-X * £10% 




Fi K . 54 

b. 7 he American Water-Bach Ore-Hearth} — Figures 2 54 and 
55 show the larger-sized furnace with three tuyere-holes m at the 
back, where the older, smaller form had only one. The hearth-box 
e (filled with lead, the charge of ore, and charcoal floating on top) 
is here also set in brickwork n. It holds about 2,500 pounds of 
lead. The work-stone g, leading to the kettle //, forms ;i separate 
easting from the hearth-box. The three sides of the furnace are 
formed by a water-cooled cast-iron jacket, ec, l-J inches thick, called 
tuyere-plate, resting on the hearth-box. The water enters at /, and 

1 Williams : "Industrial Report," " Geological Survey of Missouri*" 
1877, p. 86.— Trans. A. I. M. E., v., p. 884, 

8 Broadhead : "Geological Survey of Missouri," 1873-1871, p, 482. 



114 



METALLURGY OF LEAD. 



passes out at k. At the back of the tuyere-plate is the wind-box 
b. The blast enters this at a, and passes through three wrought- 
iron tuyere nozzles d (from 1 to 1^ inches in diameter) in tuyere 
holes m, into the hearth, at from 1 to 3 inches above the level of 
the lead. The hood placed over the furnace to carry off the fumes 
and gases is not shown in the drawings. 

The work in the American ore-hearth is continuous, as distin- 
guished from the Scotch hearth. This is made possible by water- 
cooling the sides of the furnace which protect the castings and the 
tuyeres. The fuel used is wood, charcoal, and bituminous coal. 



PLAN, 









I 






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i 
i 


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jj 






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In the ore-hearth used formerly at Rossie, 1 N. Y., and for some 
time, also, in the Mississippi Valley, but entirely abolished now, 
this cooling was effected by letting the blast circulate instead of 
the water. The resulting hot blast caused much volatilization of 
lead. 

c. TJie Mofet Ore-Hearth 2 (Figures 56 to 59).— In Fig- 
ures 56 and 57 the entire furnace is seen to rest on four pillars. 



1 Percy, "Lead," p. 289. 

2 Dewey, Trans. A. I. M. E., xviii. p. 674; Clerc, Engineering and 
Mining Journal, July 4, 1885 ; Ramsay, Scientific American Supplement, 
May 14, 1887, No. 593.' 



SMELTING IN THE ORE-HEARTH. 



115 



The hearth-box (lead basin) is not set in brickwork, and the lead 
is thus kept cool. Two furnaces are set back to back, the fumes pass- 
ing off under one hood. The lead runs through a separate spout near 
the top of the work-stone into a cast-iron kettle 31 inches in diam- 
eter and 44 inches deep, not shown in the drawings. It is set in a 



Fig. 56 




Fig. 57. 



Air 
Box 



b: 



Front View 






li'i'i ill; \\'\ i; ii ill: iiii . . 



Air 

Box 



B 



Blast -pipe 

for 
one fire 



Fig. 59. 



Section, Bhowing Tuyeres 



Air Box (inverted) 



Scule M inch — 1 foot 



oast-iron eylinder and heated from below, one cord of wood a week 
being required for the purpose. The two furnaces work inde- 
pendently. On the bottom of the lead basin rests the partition- 
box, having an opening near the bottom for the lead to enter. It 

serves as a support for the water-box, which cools the hottesl part. 
of the furnace, and upon which rests the air-box, consisting of two 
separate chambers (Figure 50), where the heated blast passes down 



116 



METALLURGY OF LEAD. 



'through the water-box by means of fourteen 1-inch copper tuyere 
pipes, seven on either side. The working opening is 15 inches high. 
A No. 5 Baker blower furnishes the blast. The fuel used is bi- 
tuminous coal. The reasons for the hot blast will be given in § 51. 
The tools, a long handled, round-pointed shovel, a round 
bar, a paddle, and a square bar are shown in Figures 60 to 64. "With 

Fig. 60. Fig. 61. Fig. 62. Fig. 63. Fig. 64. 



1 Round 



A 



«-: 



er 



1 Round 



\% Square 



-10— 



the Scotch 1 ore-hearth the round-pointed shovel is used at some 
works for ore only, and a square-pointed one for working in the 
furnace ; then a short scraper is introduced through the feed -door 
to remove slag adhering to the tuyere nozzle. 

§ 49. Mode of Working in the Ore-Hearth.— A fire is 

first kindled with wood ; then coal is added, and the blast started. 
This will soon set the fuel in a blaze ; more coal is then added, 
ashes and clinkers are removed, and in a short time a body of 
glowing fuel is obtained, filling the entire hearth. Some residue 
from a previous run is then spread over the back part of the fire 

1 Percy, "Lead," p. 282. 






SMELTING IN THE ORE-HEARTH. 117 

and the first charge, from ten to twenty pounds of ore, given. This 
soon becomes red-hot, and the lead that is set free trickles through 
the body of the fuel and collects on the bottom. More ore is added. 
The contents of the hearth are then lifted up with the bar and 
kept open in order that the heat may be distributed through the 
entire mass. Parts that have become melted and form lumps are 
drawn out with the shovel on to the work-stone, and the slag is sep- 
arated from rich residue, which is returned to the furnace. Ore 
and fuel are again added, and the operations continued until the 
hearth is full of lead. 

The hearth is now in normal working order ; it is filled with 
lead, and glowing fuel, mixed with partly fused and partly re- 
duced ore, floats on top of it. From twelve to thirty pounds of 
ore are mixed with 1-J or 2 per cent, of lime, spread over the 
glowing floating mass, and then covered with a little fuel. It 
is now exposed for from three to five minutes to the action of heat 
and blast. After this one of the furnace-men inserts the bar at 
different places into the lead, loosens and stirs the charge, and 
raises it slowly, for which purpose the paddle is sometimes used. 
The other man draws the semi-fused mass out from below with 
the shovel upon the work-stone; this allows the ore on the surface 
to sink down. What has been drawn upon the work-stone is broken 
up and the slag separated from the half-decomposed ore. The 
former is thrown aside (sometimes into a water-box); the latter 
goes back to the furnace, slacked lime being spread over it, if 
necessary. Any slag adhering to the tuyeres is then removed ;. 
some fuel is distributed in front of them and over the charge; 
fresh ore is spread on the fuel, and this again covered with fine 
fuel; then all is ready for a second operation. As the smelting 
proceeds the bulk of the lead that is set free trickles through the 
charge into the hearth-box and overflows through the groove in 
the work-stone into the kettle. Here it is sometimes poled before 
being ladled, By phoned, or drawn off through a spout into moulds. 
Some lead passes off with the fumes and the rest goes into the 
si air. 

Pattinson 1 calls attention to the following points in managing 

the ore-hearth. The amount of blast and its distribution through 
the entire charge should be carefully regulated, the half-reduced 
Ore should be exposed on the work-stone to the oxidizing action of 
the air, and the additions of lime and fuel judiciously made 

1 Percy, "Lead," p. 288. 



118 



METALLURGY OF LEAD. 



Two men work as partners in eight-hour shifts. The results ob- 
tained with different furnaces are given in the following table : 



•aouajsjan 


rH«»T(( lO 


a 
g 

a 
p 

g 

8 

a 


•sjaqstiq 


*2.5 
5.40 lbs. 


•spaoo 
'poo^ 


:^ : : : 


•sjaqsnq 
'I'Booa'BqO 


. . CO CO 


•spusnq 


CO • • • • 

T— 1 ' . . 


•sqi 

'I'BOO 

•rarma 


374 
2,160 


•&mo 


q fZ ui tI8J\[ 


50^ CO CO CO 


to 

I 

fa 


•aao 
nioaj }09j 

-ip P-B8I JO 
•^U80 .18(1 


74.44 
73.88 
83.90? 
73.20 

45.00 


•sqi 
'sanoq ^g 

Sv.\S AvaQ 


602 
5,247 


sqi 
l s.moq fg 
ui p-eai 


^050-* CO 
CD 00 O CO CO 
OCO^IOtj^ o 

ooi>j>t-T co 


'satio 


•sqi 
q jg m o jo 


CO 00 o o o 
10X00 o 
i>OSOO O 
© C5 C5 CD i> 
i-i CO 


h 

o 

p 


S9J8Xli; 

jo 'ON 


HiHBJrH £- 


to 

55 


Cast-iron 

Air-jacket .... 
Water-jacket . 
Water-jacket.. 
Water and air 
iacket 




•q^daa 


CO CO ^-i CO o 


•q;P!M. 


CO CO CO CO r* 


•jp'eq 

O} 1UOJ.J 


CO^ ^ • CO 

co co 5^ • co 

CO " 


i 


Keld Head Min'g 

Co., England . . 

Rossie Works, NY. 

Granby, Mo 

Hopewell, Mo 

Lone Elm, Mo. . . . 



s 



§ * 

■s « 

CO 
S3 

1- £ 

o — 

« j 

c4 '8 



3 ^ 

CO Ttl 

a ~ 

o ^ 

o . 

-H & 

cu ,. 

3 "• 

«— -*j 

_g O 

'g ^ 



S.S 



tt 



T^ 


s- 


03 


a 


13 


a> 


T3 


.4 


Ctj 


o 


,0 





o 


>> 


■fi 




a 


CU 


-i 


> 


&-, 


ea 


O 



a <v 



O -*3 

Sji 

O u 

o 
S > 



03 

* g 

0) o 



SMELTING IN THE ORE-HEARTH. 119 

The composition of some Missouri leads smelted in the ore- 
hearth is as follows: 

GRANBY. 1 HOPEWELL. 3 LONE ELM. 3 

As 0.00124 0.00583 0.00011 

Sb 0.01085 0.00803 0.00146 

Ag 0.00057 0.00219 0.00056 

Cu 0.00780 0.00585 0.01782 

Ni 0.00087 trace 0.00077 

Co 0.00005 

Fe 0.00367 0.00145 0.00686 

Zn trace 0.00156 0.00033 

S trace 

Pb 99.96905 99.97509 99.97204 

by difference. 

100.00000 100.00000 100.00000 

1. William's " Report,' 1 p. 63 ; 2. Trans. A. I. M. E., v., p. 326 ; 3. Trans. A. I. M. E. y 
xviii. p. 687. 

§ 50. Treatment of Slags. 1 — The gray slags obtained in 
smelting lead ores in the ore-hearth consist of lumps of a more or 
less scorified mixture of gangue and various lead (zinc) compounds, 
which contains mechanically enclosed lead and often particles of 
fuel. The mechanical analysis of Lone Elm slag' J gave 21.45 per 
cent, metallic lead and the remaining 78.55 per cent, of pulp 
showed the following composition: 

Residue. 

Per cent. Per cent. 

SiOo 1.97 

PbS0 4 0.24 

Fe 2 3 1.67 

ALO., 0.21 

ZnO 0.57 4.66 

PbS0 4 4.94 

Acetic Acid Solution. 

SiO, 10.73 

PbO 33.55 

Fe g 3 1.23 

A1 2 8 0.57 

ZnO 13.96 

CaO 11.49 

MgO 0.12 71.65 

Nitric Acid Solution. 

PbS 14.73 

FeS 2 0.(17 

ZnS 3.64 19.04 

100.39 

1 Bergen, in Commissioner Raymond's "Report,"' 1875, p. 1-1. 

2 Trans. A. I. M. E., xviii., p. 0H5. 



120 METALLURGY OF LEAD. 

The gray slag is smelted in a low rectangular blast-furnace, 
usually 4 feet in height, having an external crucible (the cast-iron 
lead-pot) and one tuyere at the back. The bed-plate, sloping from 
back to front, carries the fire-brick lining of the furnace, and over- 
laps the lead-pot. A brasque bottom, consisting of equal volumes 
of clay and coke, is tamped on it, 5 inches at the back narrow- 
ing down to 1 inch at the front. The front of the furnace has a 
cast-iron plate with an opening near the bottom. This is closed 
by ramming in a breast of clay over a wooden plug. The lead-pot 
is divided into two unequal parts by a partition descending nearly 
to the bottom. The melted charge, "black slag," flowing through 
the opening made by the wooden plug, passes over the larger divis- 
ion, filled with charcoal, into a tank, through which a slow stream 
of water flows. The lead filters through the charcoal, collects at 
the bottom, and is removed at intervals from the smaller division. 
A hood serves to carry off the fumes. The dimensions of the 
furnace are: 

Inches. 
Width of hearth 26 

Depth, front to back 36 

Height 46 

Diameter of tuyere 3 

Height of tuyere above bed-plate 10 

Height of pozt 9 

Width of pozt 12 

Prolonged axis of tuyere, strikes front lining of fur- 
nace above pozt 2 

Slope of bed-plate to 1 foot 1% 

Size of bed-plate 40 x 48 

Size of front plate 30 x 36 

Long diameter of lead-pot 36 

Front to back of lead-pot 12 

Depth of lead-pot 12 

Thickness of lining, back and sides 9 

Thickness of lining, front 5 

The starting of the furnace is very simple. A wood fire is 
kindled on the bottom, the furnace filled with clean charcoal, and 
some blast let on. When the charcoal is thoroughly ignited, a 6- 
inch bed of coke is placed on top and full blast (9 ounces) given. 
Clean slag of moderate fineness is then charged with the necessary 
coke, four shovels of slag to one of coke. To save the walls of the 
furnace from being eaten out too quickly, the slag is placed rather 
around the back and the sides of the furnace, and the coke in the 
front and in the centre. The wooden plug in the clay breast is 



Fig. 65. 




122 METALLURGY OF LEAD. 

withdrawn half or three-quarters of an hour after the full blast is 
given, when the "black slag" that had collected will begin to flow. 
During the run the charcoal in the lead-pot is removed every two 
or three hours. After smelting about sixteen hours the lining has 
become corroded, and infusible masses have accumulated on the 
bottom. The charging is then stopped and the furnace blown out 
by running off all the slag, tearing out the breast, drawing the 
rest of the charge, and chilling with water. After allowing 
it to cool for eight hours, adhering crusts are removed with steel 
bars, the furnace is patched, a.id it is then ready for another 
run. 4 

From fifteen to Eighteen and a half tons of slag, assaying from 
35 to 40 per cent, of lead are smelted in sixteen hours, giving 
7,500 pounds of hard lead. 

Two furnace-men, four helpers, and one roustabout form the 
crew; the fuel consumed is two tons of coke and twenty-two bush- 
els of charcoal, besides that for the boiler. 

A more modern slag-eye furnace, shown in Figures 65 to 
67, is the one used at the Lone Elm works. It is in many 
respects similar to the one just described. The furnace has eleven 
tuyeres of 1-^ inches in diameter, passing through water-boxes; 
seven of these are below the charging-door and four above the 
bottom, the blast being heated. For the reason of this see next 
paragraph. The tools used are three 6-foot bars, three shorter 
ones, two shovels, and a ladle. A furnace runs from fifteen to 
twenty days without repairing. In the summer of 1891 the brick 
walls were replaced by water-jackets, which greatly prolongs 
the life of the furnace. The composition ■ of the final " black slag " 
is: 

Si0 2 42.10 

PbO 25.37 

FeO 7.91 

A1 2 3 9.58 

(Ni,Co)0 traces 

MnO 0.27 

ZnO 4.48 

CaO 7.97 

MgO 1.66 

S 0.22 

99.56 
1 Trans. A. I. M. L., xviii., p. 694. 




ft 



Fig. 66. 



I . I 



m 



HE 



^-2-4 



i r 



— h — ■ 



o 



EH 



T~ il 



m 



T~^l 



Blast ripe 



P©J 



EEJ 



i — r 



mnn>mM/)M,7 7x re I i 1 I I 



1 T 



w///)//j////wm 



Elevation of the back 



mmrmmrnTmam 



r 



124 



METALLURGY OF LEAD. 



§ 51. Recovery of Fluedust by the Lewis and Bartlett 

Bag Process. 1 — One of the principal disadvantages of the hearth 
treatment is the loss in lead by volatilization. In some instances ~ 
condensing apparatus are used to save the flue dust. This subject 
will be discussed in connection with the smelting of argentiferous 
lead ores in the blast-furnace (§ 85). With most American ore- 
hearths the fumes are allowed to go to waste. The Lone Elm 
works began as early as 18/6 to collect and utilize them for pig- 
ments. This has been carried so far that in smelting lead ores the 
aim is not to produce as much metallic lead as possible, but to ob- 

Fig. 67. 
^Blast-Pipe 



at 



Water Boxes 



tb 




Section through upper Tuyeres 



tain large amounts of fume, which are condensed, purified, and 
sold in the market. This explains the use of hot blast in both 
furnaces and the position of the seven upper tuyeres below the 
charging-door in the slag-eye furnace. This method of working 
permits the treatment of large amounts of ore and shows also why 

1 Dewey: Trans. A. I. M. E., xviii., p. 674 ; Clerc : Engineering and 
Mining Journal, July 4, 1885 ; Ramsay : Scientific American Supplement, 
May 14, 1887, No; 593. 

2 Roesing : Zeitschrift fiir Berg-, Hiltten- unci Salinen-Wesen in 
Preussen, xxxvi., p. 103; (Lead Smelting in England) "English Govern- 
ment Report," quoted in § 48, a. 






SMELTING IN THE ORE HEARTH. 



1'25 



only 50 per cent, of the ore charged is converted into metallic 
lead in the ore-hearth. 

The fumes are recovered by filtering them through woollen bags. 
This has been tried at the Grant smelter at Leadville, 1 but was 
given up again, as the flue dust saved, while running high in lead, 
contained only from 6 to 10 ounces silver per ton, and this did not 
pay at that time for the condensation. The process is, however, 
introduced at the Globe Smelting and Refining Works at Denver* 



Fig. 68. 



Hot Blast Pipe 




Plan of Cooling Cylinders etc. 

and promises well. The conversion of the lead fumes into a mar- 
ketable pigment comprises two operations: 

1. Cooling and collecting the dark ore-hearth fume, "blue 
powder," in the first bag-house, the "blue room." 

2. Refining the blue powder in the slag-eye furnace, which is 
followed by cooling and collecting the resulting "white paint" in 
the second-bag-room, the "paint-house." 

1 Ghiyard, in Emmons' " Geology and Mining Industry of Leadville," 

monograph xii. in " United States Geological Survey," pp. »;?:'> and 717* 



126 



METALLURGY OF LEAD. 



The fumes from the ore-hearth are drawn off by a suction-fan, 
6 feet in diameter and 3 feet wide, which makes 290 revolutions 
per minute. They pass first through a brick dust-chamber (40 feet 
long, 19 feet high, and 6^ feet wide, with a door on one side), 
where any coarse-grained particles of more or less changed ore and 
fuel are collected. They thus pass out of the top of the chamber 
through a horizontal sheet-iron pipe, 5 feet in diameter, resting on 
20-foot iron pillars, to the fan, and thence through a 4-foot pipe, 
resting on 12 -foot pillars, to the blue room. The pipe is suf- 



Fig. 09. 




-8#-> 



,-K 



V 



O 



v 



Section of the Cooling Cylinders showing 
the heating of the blast for the Slag-Eye 
Furnace. 

ficiently long (350 feet) for the gases to cool in their passage 
through it. 

The first bag-house is similar in construction to the second one, 
shown in cross-section in Figure 71. It is a brick building (95 feet 
long, 50 feet wide, and 45 feet high) divided into two compart- 
ments by a longitudinal wall, so that one may be shut off when it 
is necessary to gain access to the bags. Each compartment is 
divided into two stories, the lower being 12 feet high. The divis- 
ions (columns, beams, etc.) are all made of iron pipe. In fact 
everything in the building is either of brick or iron, except the 



SMELTING IN THE ORE-HEARTH. 



127 



filtering bags. The lower story contains four rows of sheet-iron 
hoppers, extending the length of the building, which serve to col- 
lect the fume that has accumulated. They have the form of a 
truncated pyramid and are closed at their lower face by a sliding 
damper. They stand on four iron pipes, 3-£ feet long, encased in 
refractory clay pipes. The upper face of a hopper is covered with 
sheet iron -fa inch thick. This has 16 holes, 18 inches in diam- 
eter, from which thimbles, 12 inches high, project upwards. Over 
these the lower ends of the bags, made of unwashed wool, 60 inches 
in circumference and 33 feet long (changing to 50 inches and 35 
feet when in use), are slipped and tied fast. The upper ends are 
tied with strong cord, with which they are suspended from beams 
near the roof. There are 800 bags in the bag-house, each costing 
$9. Between every two rows of bags is an iron scaffolding with 
iron footways placed at convenient heights, so as to make all parte 
of the building accessible. 

The cooled gases, being pressed through the main pipe, enter 
four branch pipes, each of which passes through and connects a set 
of hoppers. The gases, laden with the dust, ascend into the hang- 
ing bags, where they are filtered, the fumes falling into the hop- 
pers below. These are emptied once in two days, when the bags; 
are also shaken to detach adhering fume. For this purpose the 
current of the gas is shut off, and men with •aspirators pass quickly 
through the building, giving each bag a quick shake. 

The collected fume is a very fine bluish-gray powder (blue 
powder), consisting mainly of lead oxide and sulphate, with some 
lead sulphide. 



BLUE POWDER. 



Pbs 

PbS0 4 

PbO 

Zn 

(FhA1) 8 O b 

CaO 

BtO. 

co 2 

so 2 

On combustion yield. m| 

H.O 

(X> 2 



1st Hopper, 

1st Row. 



6.18 
45. 34 
44.44 
.01 
.10 
.21 
.17 
.38 
M 



2.12 
6.68 



6th Hopper, 
2d Row. 



10.37 
43.55 
44.48 
.34 
.05 
.01 
.11 
.19 
.12 



1.57 
8.98 



1st Hopper, 
4th Row, 



5 1!) 
46.88 
45.08 
.48 
.07 
.08 
.12 
.38 
.68 



2.88 



10th Hopi>e 
4th Row. 



8.61 

43.57 

44.18 

.61 

m 

.02 
.14 

.(IS 

.44 



1.71 
2.112 



128 METALLURGY OF LEAD. 

The fume let down from the hoppers is spread over the floor 
in piles and set on fire with oil-waste. It burns for about ten 
hours and does not flame, but liberates a good deal of heat and 
some sulphurous acid. The tine, loose blue powder is hereby con- 
verted into a porous, pinkish-white crust that is friable but suffi- 
ciently coherent to stand handling and charging. The roasted blue 
powder is free from carbonaceous matter and lead sulphide. 

ROASTED BLUE POWDER. 

PbS0 4 48.76 

PbO 46.82 

Fe 2 3 0.32 

Al 1 O a 0.05 

ZuO 0.27 

CaO 0.48 

Si0 2 0.10 

CO 2 0.90 

SOi 1.65 

H 3 0.37 

99.72 

In refining the roasted blue powder in the slag-eye furnace, the 
object is to oxidize all the components of the charge as much as 
possible ; hence little metallic lead is produced. To prevent any 
carbonaceous compounds from injuring the color, Connelsville coke 
is used as fuel. 

The average daily charge for a furnace is made up of 2,800 
pounds slag, 1,000 blue powder, about 600 dry-bone (carbonate 
ore), 450 fume from the cooling pipes, and the necessary coke. 
For every pound of metallic lead about 1.6 pounds of paint are 
•obtained, and the product of a furnace is 4,250 pounds of paint. 

In running the furnace, it is important to have a hot top, a 
liquid slag, and no stoppages or irregularities, otherwise the paint 
obtained will be of inferior grade. Thus, when a furnace is 
started, the product of the first four or five hours is classed as blue 
powder, and that of the next twenty hours is still below the 
standard. 

A No. 5 Baker blower furnishes the blast for three slag-eye 
furnaces. From these the gases are drawn by a fan (6 feet in 
diameter, 3 feet wide, making 290 revolutions per minute), first 
through a set of cooling pipes which surround the blast-pipe, and 
then through a second set built over brick dust-chambers, where 



SMELTING IN THE ORE-HEARTH. 



129 



any heavy particles are collected to go back to the furnace, while 
the gases sufficiently cooled pass off into the second bag-room. 
The first set — shown in vertical section in Figure 69 and connected 



r\ 



Fig. 70. 




WHITE FUME COOLING-PIPES 

Plan and Elevation. 

Scale 1 Inch - 10 feet. 



with the second set in Figure 68 (plan) and Figure 70 (plan and 
section) — consists of two iron cylinders, 7 feet in diameter and 20 
feet high, lined with fire-brick and oonneoted by a pipe 3^ feet in 



130 METALLURGY OF LEAD. 

diameter. The blast-pipe, 18 inches in diameter at its entrance 
near the bottom, widens to 3 feet, to take up more heat from the 
surrounding gases, and before passing out is again contracted to 
its normal size. The second set, four U-shaped iron cooling pipes, 
are 3 feet in diameter and 20 feet high. 

The second bag-room or paint-house, shown in Figure 71, is a 
brick building without any partition wall. It is 40 feet wide, 90 
feet long, and 45 feet high. It has two stories ; the lower is 9 feet 
high. The divisions and general arrangements are similar to those 
in the first bag-house. The lower floor, however, is boarded. The 
hoppers are made of wood and lined with iron. They are sus- 
pended by iron straps (12 inches apart) from l|-inch pipes. These 
are laid across the 2-inch pipes at intervals of 2 feet. Beneath the 
hoppers are wooden bins closed with canvas. 

The paint from the bins is packed into barrels, each holding 500 
pounds. It has a good body, a good color, and mixes well with oil. 
It is also used as weighting material. Its price in 1884 was $3.50 
per hundred pounds, with pig-lead at $3.35. Its composition is 
shown by the following analyses : 

• Insoluble 0.08 0.08 

PbS0 4 65.46 65.00 

PbO 25.85 25.89 

ZnO 5.95 6.02 

Fe 2 3 0.03 0.03 

CaO 0.02 0.02 

C0 2 1.53 2.00 

SO a 0.04 none 

H 2 0.69 0.85 

99.65 99.89 



Fig. 71 




CHAPTEE VIII. 

SMELTING IN THE BLAST-FURNACE. 1 

§ 52. Introductory Remarks. — The treatment of lead ores 
in the blast-furnace is generally discussed under three heads : 

Precipitation, ) £ . . . , ^ . 

Roasting and Reduction, \ for S^lph^e Ores. 

General Reduction, for Oxidized Ores. 

These processes are quite distinct in European practice, where 
large mines furnish certain smelting works with uniform ores. For 
instance, in the Hartz Mountains, in Prussia, concentrated galena 
ores have been for years, and still are, smelted raw in the blast- 
furnace. In other districts of Germany as well as in France the 
sulphide ores are always roasted before they come to the blast- 
furnace, and oxidized ores are of such rare occurrence that they 
are hardly treated separately. 

In the United States this regularity of treatment can seldom be 
pursued, as the ores a smelter has to treat show the greatest vari- 
ety, as can be seen from § 35, Purchasing of Silver-Lead Ores. 
For this reason the chemical classification suited to European prac- 
tice will be set aside and smelting in the blast-furnace treated as 
one process, in which lead and silver are extracted from the ore in 
the form of base bullion by means of carbon, iron and lime. 

All smelting ores can be worked in the blast-furnace ; any ore 
containing over 4 per cent, of silica must be so treated. If it is a 
sulphide, it is generally first roasted in the reverberatory furnace ; 
if a carbonate or a mixture of carbonate and sulphide, the carbon- 
ate prevailing, it is smelted at once. If the preceding analyses of 
argentiferous lead ores of the Rocky Mountains (p. 33) and the 
Pacific (p. 40) be referred to, it will be seen that they usually con- 
tain much over 4 per cent, silica. This explains the universal use 
of the blast-furnace in the West. 

1 Hahn, " Mineral Resources of the United States," 1882, p. 325 ; also 
Engineering and Mining Journal, August 25, September 1 and 8, 1883. 
Henrich, Engineering and Mining Journal, October 3, 1883 ; July 17> 
1886. Guyard, in Emmons' "Geology and Mining Industry of Leadville, 
Col.," monograph xii., United States Geol. Survey, 1886, p. 613. 



SMELTING IN THE BLAST-FURNACE. 133 

§ 53. Lead SlagS. 1 — As the object of smelting is to separate 
by fusion the lead as metal from its ore, the other constitutents, 
the acid silica and the bases, iron, lime, etc., have to be combined 
in certain proportions to form a slag. If this does not occur 
naturally, either silica or the various bases, i.e., fluxes, will have to 
be added in the requisite quantities. 

Composition of Lead Slags. — The main slag for the lead smelt- 
er is the singulo-silicate slag, 

ii in 

2 RO+SiO^ or 2R 2 3 +3Si0 2 , 

with the oxygen ratio of bases to silica as 1 : 1. 

In practice the slags are made slightly more acid than the for- 
mula calls for. Another slag that is sometimes made is the sesqui- 
silicate 

n • in 

4RO+3Si0 2 , or 4R 2 3 + 9Si0 2 , 

with the oxygen ratio of bases to silica as 2 : 3, written some- 
times 

ii ii in hi 

(2RO+Si0 2 )+2(RO+Si0 2 ), or (2R 2 3 +3Si0 2 )+2(R 2 3 + 3Si0 2 ), 

as if consisting of a singulo- and a bi-silicate. This slag, not being 
so readily fusible as the singulo-silicate, is made, if a high temper- 
ature is desirable, e.y., if much lead sulphide is to be decomposed 
by metallic iron resulting from ferric compounds. With a readily 
fusible slag the ferric compound would be reduced only to a fer- 
rous compound and then slagged. Henri ch 2 recommends a slag 
more acid than the singulo-silicate for arsenical ores in order to 
keep the speise Liquid. 

The following tables by Balling 9 give the necessary proportions 
of silica and bases required to form singulo-silicate, bi-silicate 
and sesqui-silicate slag, one table having silica for a unit, the other 

the bases. The base baryta has been added. 

1 Eahn, Op. n't., p. 825 ; Ghiyard, <>}>. cit., p. 701 ; lies, "Mineral Re- 
sources of the United States," 1883-84, p. 440; Schneider, Trans. A. I. 
M. E., xi., p. 56. 

8 Engineering <ni<! Mining Journal, October G, 1883. 

3 "Compendium der Metallurgischen Chemie," Bonn, 1882, p, 98. 



m 



METALLURGY OF LEAD. 



One Part by Weight of 
Silica Requires 



For Singulo-Silicates 

Lime 

Baryta 

Magnesia 

Alumina 

Ferrous oxide 

Ma nganous oxide. . 

For Bi-Silicates : 

Lime 

Baryta 

Magnesia 

Alumina 

Ferrous oxide 

Manganous oxide. . 

For Sesqui-Silicates : 

Lime 

Baryta 

Magnesia 

Alumina 

Ferrous oxide 

Manganous oxide . 



Parts by 

Weight 

of Bases. 



1.86 
5.10 
1.33 
1.14 
2 40 
2.36 



0.93 
2.55 
0.66 
0.57 
1.20 
1.18 



1.24 
3.40 

0.88 
0.76 
1.60 
1 57 



One Part by Weight of 
Bases Requires 



I Parts by 

Weight 

of buica. 



For Singulo-Silicates : 

Lime 

Baryta 

Magnesia 

Alumina 

Ferrous oxide 

Manganous oxide. . . 

For Bi-Silicates : 

Lime 

Baryta 

Magnesia 

Alumina 

Ferrous oxide 

Manganous oxide. . . 

For Sesqui-Silicates : 

Lime 

Baryta 

Magnesia 

Alumina 

Ferrous oxide 

Manganous oxide. . . 



0.535 
0.196 
0.750 

0.873 
0.416 
0.422 



1 070 
392 
1.500 
1.747 

883 
0.845 



0.803 
0.294 
1.125 
1.310 
0.625 
0.633 



This table is to facilitate the study of slags that do not corre- 
spond to any of the typical slags but nevertheless give satisfactory 
results in the furnace. 

Typical slags are definite combinations of silica, iron, lime and 
sometimes alumina, the success of which has been proved and which 
therefore become representative. They fulfil the requirements 
of a good slag, which are, according to Eilers, 1 that it should not 
contain over J per cent, of lead or % ounce of silver to the ton, 
provided that the base bullion does not run higher than 300 ounces, 
nor have a density over 3.6, nor permit accretions in the hearth, 
thus keeping the lead red-hot, nor any creeping up of over-fire. 

For instance, a typical slag discovered by Eilers is formulated 
thus: 

6FeO,3Si0 2 -f2CaO.Si0 2 . 

In percentage it gives: 

30.6lSiO 2 +55.10FeO+14.29CaO = l()0. 

After deducting 10 per cent, for other ingredients of ore and 
■fluxes, such as alumina, zinc oxide, etc., that cannot be brought 



1 Engineering and Mining Journal, April 9, 1881. 



SMELTING IN THE BLAST-FURNACE. 



135 



under the head of ferrous oxide, like manganous oxide, or of lime, 
like baryta or magnesia, there remains: 

27.55SiO 2 +49.59FeO-fl2.86CaO = 90, 
which gives, in round figures, 

28SiO 2 +50FeO4-12CaO+9Al 2 O 3 , etc. = 100. 

Another slag, first brought into prominence by Eilers, has the for- 
mula, 

6Fe0.3Si0 2 +4Ca0.2Si0 2 , 

and therefore the following percentage of composition: 

31.38SiO 2 +45.18FeO+23.44CaO = 100. 

Deducting again 10 per cent., there remains 

28.24SiO 2 - r 40.66FeO+21.10CaO = 90, 

which practical experience has altered to 

30SiO 2 +40FeO+20CaO+10Al 2 O 3 , etc. = 100. 

The following slags have been thoroughly tested and are in suc- 
cessful use. The first seven (A to G) are those so designated by 
lies in his paper already quoted. 



Authority. 



lies 

lies 

Eilers 

lies 

Eilers 

Schneider. . . 

Raht 

Hahn 

Page 

Livingstone 

Hahn 

lies 

Murray 



Type. 


Si0 2 . 


Fe(Mn)0 


A* 


32 


52 


B* 


35 


45 


(J 


28 


50 


D* 


34 


34 


E 


30 


40 


F 


33 


33 


G 


35 


27 


H 


34 


50 


I 


33 


36 


J 


30 


36 


K 


36 


40 


L 


32 


33 


M 


40 


34 



Ca(Ba, Mg)0. 

6 
10 
12 
16 
20 
24 
28 
12 
16 
20 
20 
23 
26 



Total. 



90 
90 
90 

84 
90 
90 
90 
96 
85 
86 
96 
88 
100 



* No longer recommended by lies. 

These typical slags are the general guides for the lead smelter. 
lie chooses the one best suited to the character of his ore. lie 
need not, however, adhere strictly to it. Slight variations are 
always liable to occur, and need give no uneasiness, if the furnace 
runs well and the slag analysis shows a low amount, of lead and 
silver. 

The old rule of making a slag with from 80 to 34 per cent, sil- 



136 METALLURGY OF LEAD. 

ica, an equal amount of metallic iron, and from 8 to 12 per cent, 
lime is practically obsolete. 

In reviewing the table, it will be seen that the composition of 
the slags ranges between 

Si0 2 28 and 36 per cent. 

FeO 24 " 52 " " 

CaO 6 " 30 " " 

Slags have been made that run as high as 40 per cent, silica, 
for instance, 40SiO 2 , 40FeO, 10-12CaO, but they require much fuel, 
and are slightly viscid even then. The lowest practical limit in 
silica is probably reached with 28 per cent. The percentages of 
ferrous oxide represent also extremes. As regards the lime, 30 per 
cent, is the highest Schneider 1 gives in his experiments, but 6 per 
cent, is too low A modern American lead slag does not contain 
less than 10 per cent, of lime; European slags, as quoted by Kerl 
and Percy, run as low as 5 per cent., although the tendency there is 
towards a higher percentage. 

PHYSICAL PROPERTIES OF LEAD SLAGS. 

The fusibility of a slag depends on the percentage of silica and 
the character of the bases; the effect of the latter will be discussed 
in the next paragraph. The more fusible a slag the larger will be 
the amount of charge smelted per unit of fuel. With the corre- 
spondingly low temperature in the furnace the reducing power will 
be diminished; therefore a very readily fusible slag is not always 
desirable. 

The liquidity of a slag depends largely on its fusibility. A 
slag ought to be sufficiently liquid to allow a perfect separation of 
lead in the furnace and of matte in the slag-pot. A correctly 
composed slag, which would otherwise be liquid, becomes viscous if 
the weather suddenly becomes wet or cold; in such cases more fuel 
is required. It is difficult, even for the practised eye, to say always 
whether the viscidity comes from lack of fusibility or from a 
slight insufficiency in fuel. As a rule, singulo-silicate slags con- 
taining earthy and metallic bases solidify quickly without first 
becoming pasty. When the tap-hole has been closed, the slag, 
if good, will drop slowly into the pot, drawing a small thread as it 
leaves the spout; a slag with an excess of base will fall off quickly 
in little round drops. Most slags have certain characteristics in 

1 Trans. A. I. M. E., xi., p. 57. 



SMELTING IN THE BLAST-FURNACE. 



137 



their manner of running, which have to be studied by observation. 
In close connection with the running of the slag from the furnace 
is the manner of its rising in the slag-pot and the appearance of 
the surface when it has just solidified and is still red. Thus many 
slags show very characteristic surfaces. 

Well-composed slags have a decided tendency to crystallize. 
The centre of a cone of slag is generally more crystalline than any 
other part, because the cooling is slower. Slags that crystallize 
usually become amorphous if chilled suddenly, and crypto-crystal- 
line if not given sufficient time to develop crystals. A small per- 
centage of zinc oxide in the slag interferes greatly with the crys- 
tallization, lies thinks that the form in which a slag crystallizes 
stands in some relation to the percentage of lime it contains. He 
says (see Figure 72): 




d I' q 





g 'h 

Fig. 72. 

Slags with from :{ to 5 per cent, of lime crystallize like a. 

a'. 
b. 

V. 

d. 
e. 
9- 

1 1 1 j 1 1 i 1 1 lc 1 < > 



8 
IS 
19 
28 
26 
80 



12 

1H 

26 

27 
86 



atte 



The tonus t l'h represent crystals formed II 
slag very siliceous ores with lime alone. It is to he noted thai the 
same Blag will melt at a lower temperature when glaflSythan when 
crypto-crystalline. 



138 METALLURGY OF LEAD. 

There is some variety in the tenacity of slags. Siliceous slags 
are generally tougher than those where the base prevails. The 
more crystalline a slag the more brittle. A pot of slag may show 
brittleness in one part and toughness in another. Slags are not al- 
ways pure silicates as expressed in the chemical formula?, they 
are often mixtures of different silicates or even chemical silicates 
holding indefinite compounds of silica and base in solution. If, in 
these, crystals form, they represent the true silicate, being sur- 
rounded by a mass of an entirely different composition and 
different physical properties. This is important to remember in 
taking a sample for slag analysis (see § 81). 

The specific gravity of a slag is an important factor in its sepa- 
ration from the matte. lies gives, as extreme figures, 3.3 and 
4.16; as a common range for good slags, 3.4 and 3.6. The average 
specific gravity of one hundred determinations of good slags gave 
him 3.691, and 3.6 is accepted as the highest specific gravity a 
good slag ought to have. 

Slags do not possess to any extent the property of transmitting 
light. Single crystals are in exceptional cases transparent; some- 
times they are translucent, but generally opaque. 

Lead slags are usually black from their high percentage of iron. 
Thin pieces sometimes appear greenish-black; a large amount of 
iron will give a brownish hue. Lime produces a bluish or gray- 
ish tone. 

The lustre of slags varies. It is occasionally metallic, but gen- 
erally vitreous to resinous. Very often slags are dull. 

Last may be mentioned the magnetic property of some lead 
slags, to which lies first called attention. This is caused by the 
presence of magnetite 1 or magnetic sulphide of iron (Fe 8 S 9 ). How 
the magnetic oxide gets into the slag is a matter for further inves- 
tigation. Hahn suggests an incomplete reduction of ferric oxide; 
Guyard thinks it results from the oxidation 2 of metallic iron by 
lead oxide: 

4PbO+Fe 3 = Fe 3 4 +4Pb. 

§ 54. Action of Fluxes and Influence of Foreign Mat- 
ter. — The fluxes usually required are iron ore, manganese ore, lime- 
stone, dolomite, rarely fluorspar, and slag. 

a. Iron. — An iron flux acts in three different ways. It gives a 
base for the silica in the ore, 

1 Guyard, Op. cit., p. 702. 

2 Berthier, " Traite des essais par la voie seche," Liege, 1847, i., p. 354. 



SMELTING IN THE BLAST-FURNACE. 



139 



FeO.Si0 2 -fFeO = 2FeO.Si0 2 , 
4FeO.Si0 8 +2Fe0.2Si0 2 =6Fe0.3Si0 8 . 

Being reduced to metallic iron by means of carbon, it acts as a 
precipitating agent (§ 7), 

4PbS+2Fe 8 3 +3C=4Pb+4FeS+3CO g , 
2PbS+4FeO.Si0 2 +2C=2Pb+2FeS+2Fe0 2 Si0 2 +2CO. 

It liberates lead oxide from its combination Avith silica, after 
which it is then reduced by means of carbon, 

4FeO.Si0 lg +2PbO.SiO g +C = 2Pb+4FeO. 2Si0 2 +C0 2 . 

The larger the amount of iron, within reasonable limits, a slag 
contains the greater will be its fusibility and liquidity and the 
higher its specific gravity. Slags rich in iron are dangerous, as 
they are liable to cause the formation of crusts in the furnace. A 
slag high in iron is, however, a necessity if much zinc is contained 
in the ore, as iron favors the slagging of zinc oxide. The purer 
an iron ore the greater will be its fluxing power, as only that part 
of the iron is available which is not required by its own silica, sul- 
phur or arsenic. The silica not only limits the fluxing power, but 
also consumes limestone to form the slag. With the increase of 
slag in the blast-furnace the amount of lead ore of course dimin- 
ishes. 

These remarks are equally applicable mutatis mutandis for 
the other fluxes. 

ANALYSES OF IRON FLUXES. 





Bruce Iron 
Ore, Col. 


Madonna 
Mine, Col. 




Bruce Iron 
Ore, Col. 


M adonna 
Mine, Col. 


Fe 


66.44 42.05 


SiO., 

CaO" 


2.39 
0.12 
0.62 
0.04 
0.18 
0.06 

1. 


5.10 
2.0-5.0 

2 ' 


Mn 

Zn 


0.01 
0.02 
0.02 

0.01 
l. 


3 12 
1.86 

4.0-9.0 

6.02 

2. 


MgO 


( u 


A1 2 3 


Pb 


Ag,oz. p. t 

Au,oz. p. t. . . . 

Reference 


As 


S 

Reference 



i Guyard, <)p. cit., p. 647. 

9 Dewey, Bulletin No. 42, " United States National Museum," p. 47. 

b. Manganese^ —The fluxing properties of manganese are sim- 
ilar to those of iron, and it may be substituted for the latter it' de- 

1 lies : " School of Mines Quarterly,'' v., p. 217, and Engineering and 
Mining Journal, March 12, 1881 ; Eilers : Engineering and Mining Jour- 
nal, April 9, 1881. 



140 METALLURGY OF LEAD. 

sired. It makes a slag more fusible and liquid, so that with large 
amounts of manganese it is advisable to choose a slag that does 
not require much metallic base. The commonest manganese min- 
eral is pyrolusite, Mn0 2 . Its oxidizing power has been said to be 
the cause that certain slags rich in manganese have so high a 
tenor in silver. This seems to be contradicted by the fact that 
slags rich in manganese have been made by Church x that ran low 
in silver and lead, the base bullion averaging 314 ounces to the ton. 



SiOo 


FeO 


MnO 


CaO 


MgO 


A1 2 3 


Ag. ozs. 


Pb 


29.60 


11.56 


43.25 


7.50 


trace 


6.34 


trace 


1.4 


33.00 


14.22 


25.78 


13.10 


1.00 


4.20 


0.5 


1.0 



The oxidizing power of oxides of manganese on blende in the 
reverberatory furnace, when sulphide copper ores are smelted, is 
mentioned by Pearce, 2 who obtained a slag of SiOa, 48; MnO, 30; 
and ZnO, 12.5 per cent.; some manganese (3 per cent.) also enter- 
ing the matte (50 to 60 percent, copper) as sulphide. lies believes 
that in the blast-furnace the amount of matte and speise formed 
diminishes with the increase of manganese in the charge. 

Harbordt 3 gives it as his experience that the percentage of 
matte formed is not affected unless the manganese is present in 
considerable quantities. Another peculiarity of manganese is that 
it reduces the dissolving power of a slag for zinc oxide, magnesia, 
and barium sulphide. 

According to Chisolm, 4 60,000 tons of manganiferous iron ores 
were smelted in Colorado in 1887. 

ANALYSES OF MANGANESE ORES AND MANGANIFEROUS IRON ORES. 

Montana. 5 Colorado. 6 

Si0 2 ... 6.60 8-15 

Fe 2 3 3.20 35-50 

Mn 2 3 88.40 8-15 

Ag.oz 9.00 7-12 

c. Lime and Magnesia. — The manner in which lime replaces 

1 Trans. A. I. M. E., xv., p. 612; School of Mines Quarterly, v., p. 322. 

2 Trans. A. J. M. E., xi., p. 59. 

3 Private communication, July, 1891. 

4 "Mineral Resources of the United States," 1887, p. 153. 

8 Peters : "Mineral Resources of the United States," 1883-84, p. 380. 
« Chisolm : Ibid. 1885, p. 348. 



SMELTING IN THE BLAST-FURNACE. 14 L 

iron in a slag can be expressed by changing the formula given 
above to 

4Fe0.2Si0 8 +2PbS+2CaO+2C = 2Pb+2FeS+(2CaO.Si0 2 + 
2FeO.Si0 2 )+2CO. 

Burnt lime is rarely, if ever, used in the blast-furnace. The 
effect of lime is to decrease the fusibility and the specific gravity 
of slags. Slags rich in lime require more heat, i.e., more fuel, a 
stronger blast, and consequently a higher furnace than if iron pre- 
dominates. They generally give a good separation of slag and 
matte, and limestone being in nearly all cases cheaper than iron 
ore, the tendency in most large smelting works is to make slags 
rich in lime. 

Schneider 1 found that with slags containing much lime less 
matte is formed than when they are rich in iron, and also that the 
matte is lower in lead and higher in silver. He explains it by say- 
ing that calcium sulphide is formed and then dissolved by the 
slag. Its presence in Leadville slags has been proved by Guyard. 2 
The use of lime in slags is limited by foreign matter in the ore, 
and especially by the presence of zinc. In a general way it 
may be said that the less lime and the more iron a slag contains the 
better will the furnace work. It does not seem advisable to go 
beyond 16 per cent, of lime in a slag, if (say) from 9 to 12 per 
cent, of zinc is present in the charge. With 28 per cent, of lime 
the zinc simply refuses to enter the slag; it is volatilized and in- 
crusts the furnace. 

Magnesia is undesirable in a lead furnace, as it makes a slag 
pasty and streaky, but in many cases the only available limestone 
is dolomitic, and it must be made the best of. This undesirable 
property of magnesia is especially observable if the slag contains 
zinc. Magnesia and zinc- oxide appear to intensify each other's 
property of being difficult to slag. In a slag containing 8 per 
cent. /.in<- and from 2 to 3 per cent, baryta, very common just now 
in Colorado, from 2 to 3 per cent, magnesia shows a decidedly bad 
effect, and 5 per cent, causes a great deal of trouble. 

Magnesia is generally figured in a slag as replacing lime; it 
may be doubted whether this is justifiable (Harbordt 8 ). 

1 Trans. A. I. M. E., xi., p. 58. 

8 Op. cit., p. 73. 

■ Private communication, July, 1891. 



142 



METALLURGY OF LEAD. 



ANALYSES OF LIMESTONE. 



CaC0 3 .. 
MgCO,... 

Si0 2 

(FeAl) 2 0, 
FeCO q . .' 



Reference. 



Cafion 


Glass Mine, 


Carbonate 


Iron Co., 


Saint Jo- 


Mine La 


City, Col. 


Col. 


Mine, Col. 


Mo. 


seph, Mo. 


Motte, Mo. 


88.90 


49.57 


57.95 


47.50 


60.34 


55.2 


6.30 


37.08 


39.65 


42.19 


32.17 


40.9 


3.10 


4.22 


0.76 


5.11 


3.77 


1.27 


1.50 


6.23 




F 2 3 4.67 


6.88 




1. 


1. 


1. 


2. 


3. 


4. 



1, Guyard : Emmons, Op. cit., p. 646 ; 2, Gage, Trans. A. I. M. E. s iii., p. 117 ; 3, Monell, 
School of Mines Quarterly, ix., p. 214 ; 4, Neill, School of Mines Quarterly, ix., p. 214. 

d. Fluorspar is of slight importance in smelting lead-silver ores 
in the blast-furnace. It forms with barium and calcium sulphate 
readily fusible compounds, and assists in fluxing zinc either as sul- 
phide or oxide. This is mainly due to the fact that fluorspar, 
when once melted, is very liquid, and thus assists other less fusible 
compounds to enter the slag. Its chemical action in volatilizing 
silicon as silicon fluoride need not be taken into consideration. 
Foehr l makes a great many claims for fluorsjw: that the use of 
from 1 to 5 percent, in roasting ores in a reverberatory furnace 
saves fuel, that adding it to the charge in refining lead prevents 
shots of lead from being retained by the litharge, etc. 

e. Slag. — There are four reasons for the use of slag in a blast- 
furnace charge. (1) It may contain too much lead or silver to be 
thrown away. (2) It makes the charge less dense. (3) It helps the 
actual smelting process because the slag having been already melted, 
it will re-melt easily and promote the smelting of the ore itself; 
and (4), if it be more acid or basic than the slag that is being 
formed by the smelting mixture, it will act as an acid or basic flux. 

With a furnace running in a normal way, some rich slag is al- 
ways produced. This is especially the case when the last slag in 
the furnace is being tapped and the blast passing through the tap- 
hole blows out valuable parts, which enrich the slag in the pot 
(blow-pot). Then again, when much matte comes out with the 
slag, it is liable not to settle out perfectly. 

With a coarse easy charge the addition of slag is not absolutely 
necessary, although some is always given — say from 150 to 200 
pounds for 1,000 pounds of ore. With fine ore the addition of slag 
is necessary, as otherwise the blast would not penetrate evenly 



Engineering and Mining Journal, June 21, 28, 1890. 



SMELTING IN THE BLAST-FURNACE. 143 

through the smelting mixture, but form blow-holes. As much as 
25 per cent, of the charge may have to consist of slag. With ores 
rich in zinc, slag is added to increase the fusibility. 

Slag more basic than the normal slag comes into play where, 
for example, matte is being concentrated in a reverberatory fur- 
nace, and the resulting slag contains much iron that is available for 
the blast-furnace. 

In smelting the by-products of desilverizing works in the blast- 
furnace without ore, the amount of slag added goes up as high as 
50 per cent. 

f. Alumina. — From the composition of the typical lead slags it 
will be seen that the place alumina occupies in lead smelting is 
generally a subordinate o e. When it is present in large quanti- 
ties, it becomes a question whether it acts as an acid or a base. It 
is known in a general way that with a high percentage of silica, 
alumina acts as a base ; with a low percentage it acts as an acid, 
lies 1 gives this as his experience in lead smelting. Hahn, 2 how- 
ever, thinks that alumina always acts as a base, and says that an 
increase of alumina requires also an increase of silica, or, what 
would be the same, a decrease in the bases. 

Schneider 3 found that as a general rule an increase of alumina 
called for an increase in the proportion of lime, which means a de- 
crease in silica, the alumina acting as an acid. Howe,* summing up 
the statements of Hahn and Schneider, suggests that the part played 
by alumina may depend upon the proportion of the other two 
fluxes, lime and iron, and that in calcareous singulo-silicates, low 
in iron, alumina may act as an acid and in ferruginous slags, low in 
lime, as a base. The idea seems to be confirmed by the experience 
of Peters 5 in smelting Mount Lincoln 6 ores in Colorado. The 
writer's own impression is that alumina acts as an acid in ferrugi- 
nous slags, as in smelting an ore resembling an argentiferous clay lie 
could not get his furnace to work with a slag that ran higher than 
28 per cent, silica. 

Henrich, 7 in a very interesting paper on smelting sulphide copper 

1 " Mineral Resources of the U. S.," 1883-84, p. 433. 

2 " Mineral Resources of the U. tt.," 1882, p. 328. 
8 Trans. A. I. M. E., xi., p. 57. 

4 Engineering and Mining Journal, November 17, 1883. 

6 Engineering and Mining Journal, November 24, 1883. 
8 Trans. A. I. M. K, ii., p. 310. 

7 Engineering and Mining Journal, July 17, August 26, October 2, 
1886 ; December 27, 1890. 



144 METALLURGY OF LEAD. 

ores rich in silica, alumina, magnesia, and low in iron, came to the 
conclusion that alumina always acts as an acid, and recommends 
two general types of silicate-aluminates for ores rich in alumina, 
on the hypothesis that 2 A1 2 3 are equivalent to 3Si0 2 . 

12 oxygen of base : 12 oxygen of acid =1:1. 

^"•I^hfo^l = {SSfc* [Bi-Silicate-Alu^ate. 
12 oxygen of base : 24 oxygen of acid = 1:2. 
The theoretical composition of these two types is : 

\ I. n. 

Si0 2 16.1 35.0 

A1 2 3 18.2 13.2 

FeO 25.7 37.3 

CaO 40.0 14.5 

By a combination of them he thinks it possible to meet all cases 
where the percentage of alumina is high. Stone 1 and Elbers 2 have 
discussed the matter from the iron-metallurgist's point of view. 

To sum up: the question, under what conditions alumina ceases 
to be a base and becomes an acid, is not settled. It has to be 
determined for each slag by synthetical experiments in the labora- 
tory. 

g. JBarite. — Barium sulphate in the blast-furnace is partly con- 
verted into barium sulphide and partly into barium silicate. In 
both forms it causes an imperfect separation of matte and slag. 
The reaction is usually expressed by the following formula: 3 
2BaS0 4 +8Fe+4Si0 2 =(BaS.FeS) + 7FeO.Ba0.4Si0 2 . 

This would account for the formation of barium sulphide, even 
if no barium sulphate were reduced by means of carbon. It would 
also lead one to suppose that all barium sulphide entered the 
matte, but the fact is that, while a very little can be discovered in 
the matte, most of it is dissolved hj the slag. Further, the pro- 
portions given do not correspond to the results obtained in a lead- 
furnace. That barium sulphide can enter the matte under suitable 
conditions is proved by the mattes made in the Altai mountains, 
which contain, according to Jossa and Kurnakoff, 4 from 4 to 22 
per cent, barium. 

1 Engineering and Mining Journal, November 24, 1883. 

8 Engineering and Mining Journal, March 27, July 31, 1886. 

9 Balling, " Metallurgische Chemie," p. 89. 

* Berg- und Huttenmannische Zeitung, 1886, p. 547. 



SMELTING IN THE BLAST-FURNACE. 145 

According to Schweder 1 there is another and more important 
reaction : 

3BaS0 4 +4Si0 2 +FeS = 3BaSi0 3 +FeSi0 3 H-4S0 2 
BaSO i +2Si0 2 +Fe=BaSi0 3 +FeSi0 3 +S0 2 . 

This means that in the presence of a metallic sulphide or a 
metal, barium sulphate is readily decomposed by silica; the liber- 
ated sulphur trioxide is decomposed at the elevated temperature 
into sulphur dioxide and oxygen, and this has an oxidizing effect on 
the iron sulphide or the metallic iron. 

Schweder found in his experiments that calcium sulphate and 
sodium sulphate act in the same way. 

Silicates of iron and baryta form very liquid slags. 

In computing a charge the baryta is figured as replacing part 
of the lime. 

From what has been said, it will be clear that all the sulphur 
contained in the barite cannot be figured into the charge. In fact, 
if 19 per cent, of the sulphur is figured as combining with iron to 
form iron sulphide, the amount will be covered. 

h. Pyrite. — If an ore containing pyrite be smelted directly in 
the blast-furnace, it will consume iron as shown by the formula: 

FeS 2 +Fe = 2FeS. 

In calculating a charge this iron has to be added. There are, 
however, cases in smelting oxidized ores when part of the sulphur 
pusses off as sulphurous acid and comparatively little matte is 
made. Henrich * gives his successful experience in smelting car- 
bonate ores at Benson, Ar., consisting of galena (15 percent.) and 
anglesite (75 per cent.), with silver-bearing pyrite. He obtained 
very little matte (20 pounds from 13 or 14 tons of ore), but con- 
siderable sulphurous acid. This he explains as having been caused 
by the following reactions: 

2FeS 2 +5PbSO 4 +SiO 2 =5Pb+Fe 2 SiO 4 -f0SO 2 . 
PbS0 4 +PbS = 2Pb+2S0 2 . 

He ^ays that the furnace ran rapidly and became very hot, so 
that the fuel (coke) had to be cut down from 12.5 to 1 I per cent., 
the pressure of the blast being I | inelies mercury. 

i. ChalcopyriU <m<i ('<,/,/>,,- Ores. — In the blast-furnace the 

aim is always to carry the copper into the matte, which it enters as 

1 Berg- v. Buttenm. Ztg., 1870, p. 88. Iron, 1879, xiii., p. 387. 
* Engine*' ri in 1 and Mining Journal, September 22, lbb3. 



146 METALLURGY OF LEAD. 

cuprous sulphide. Copper, having a greater affinity than any- 
other metal for sulphur, will generally take up all the sulphur in 
the charge to form Cu 2 S, and what is left is then available for 
iron, lead, etc. If a charge does contain copper and not enough 
sulphur to form cuprous sulphide, some copper will be reduced to 
metal and be alloyed with the lead. The alloy hardens, sinks to 
the bottom, and closes up the passage of the lead-well. There is 
one case where even with sufficient sulphur to form cuprous sul- 
phide the copper combines with the lead. This is when matte is 
concentrated in the blast-furnace with a highly ferruginous slag. 
The affinities of copper and iron for sulphur and silica seem to be- 
come disturbed. 

If the slag be too basic the iron takes up some sulphur and goes 
into the matte instead of separating out and forming a crust; 
thus some sulphur belonging to the copper may be taken away, 
and this alloys with lead. Another way of explaining the fact 
would be that reactions between sulphides and oxides of copper 
take place similar to those of the reverberatory furnace, and the 
resulting metallic copper becomes alloyed with reduced lead. 
Whatever may be the theory, the fact remains that any excess of 
iron has to be avoided in the slag if the copper is to be concen- 
trated in the matte and not partly driven into the lead. This 
is liable to occur when the matte contains about 12 per cent, of 
copper, and increasingly so with the increase of copper. 

j. Blende and Zinc Oxide. — Zinc is the metal that causes the 
greatest difficulty in the blast-furnace in whatever form it may 
occur. Blende l is decomposed by iron oxides and silicates, the re- 
sulting zinc oxide entering the slag; metallic iron liberates metal- 
lic zinc, but most of the zinc sulphide entering the blast-furnace 
remains un decomposed and enters the matte as well as the slag. 
Generally the percentage of zinc found by analysis in matte and 
slag will be about equal. Blende makes matte less fusible, ob- 
structs the separation, and carries other metallic sulphides into the 
slag. "With ores containing little zinc this imperfect separation 
can be remedied by an addition of chalcopyrite 2 to the charge. 
If blende is present to any considerable extent, the ore must be 
roasted before smelting, to form zinc oxide. In order to carry this 
off, the slag has to be very fusible, as zinc silicate proper is infusi- 
ble, and thus lowers the fusibility of any other silicate. If zinc 

1 Plattner, Berg- u. Huttenm. Ztg., 1854, p. 81. 

8 Hahn, " Mineral Resources of the United States," 1882, p. 343. 



SMELTING IN THE BLAST-FURNACE. 147 

oxide is to be slagged care must be taken that it is not reduced to 
metal, hence the smelting has to be done quickly and at a low 
temperature. This requires a slag not high in silica and with a 
preponderance of iron. If, however, a slag high in lime has to be 
used, it has become the practice to figure one-half of the zinc 
(ZnO) as replacing an equivalent amount of lime (CaO) in the 
slag. The amount of lime that the slag may retain to work well in 
the furnace then depends mainly on the percentage of matte that 
is formed; the higher this is the lower must be the lime. In any 
case, in ore smelting, a slag ought not to contain more than from 6 
to 8 per cent, of zinc, as otherwise there is too great a loss in 
metal. Slags are made that contain more than 8 per cent, zinc, 
but that this is commercially good practice nobody will maintain. 
If zinc oxide be reduced to metal in the lower parts of the furnace 
by carbon, or to a slight extent by metallic iron, it becomes vola- 
tilized. The vapor ascending carries with it lead and silver, and, 
being oxidized higher up by carbonic acid or water-vapor or by 
the oxygen taken from any readily reducible oxide, carries off met- 
al as fluedust, and forms accretions on the sides of the furnace, 
which grow in thickness up to the charging-door. If zinc oxide, 
not reduced to metal on its downward course in the charge, 
Domes, with the presence of carbon, in contact with lead sulphide 
or sulphate, it is converted into sulphide. If sufficient iron is pres- 
ent it will decompose the lead sulphide and the zinc oxide will re- 
mum unchanged. 

The aim must therefore be to remove as much zinc as possible 
from the ore before it is smelted. The various attempts to do this 
have not been very successful, with the exception of mechanical 
concentration, and this with very fine-grained ores is out of place 
on account of the intimate contact of blende with galena, and with 
oxidized ores the losses in precious metal would be too great. 

A number of plans for removing zinc from ores not suited for 
concentration have been tried or suggested. Thuni ' gives u way 
of distilling the roasted ore in inclined cylindrical retorts, the zinc 

vapors being carried oil' at the upper end and the lead compounds 

and residue removed at the lower. I Jiinaghi :; proposes to work on a 

similar plan. Binon and Grandfils 3 describe a furnace with vertical 

retorts, which are charged from the topand discharged at the bottom, 

1 Berg' und HUttenmdnnische Zeitung, 1875, p. l. 
- Engineering and Mining Journal, August ~N>. issjo. 
■ Revue univereeUe des mine*\ 1879, vol. v., p. 228. 



1-43 METALLURGY OF LEAD. 

while the zinc-vapors are drawn off by a horizontal condenser near 
the charging opening. Chenhall 1 suggests a similar method. An- 
other dry method is that of Simmonet, 2 who roasts finely pulverized 
lead and silver-bearing blende with coarse-crushed limestone. Most 
of the lead and silver pass into the limestone, which has been in 
part changed to caustic lime, calcium sulphate, and sulphide, and 
can be separated from the fine zinc ore by screening. The writer 
once experimented with this method on several kinds of ore, but 
too little silver passed into the limestone to make it a success. The 
method is, however, worth trying, as it succeeded with Simmonet, 
and may again with suitable ore. 

One of the latest suggestions is by Bartlett, 3 who wants to apply 
the ore-hearth method of Joplin, Mo. (§ 48, c), in a modified form, 
and expects not to lose any silver, which is to be concentrated in a 
more or less melted mixture of matte and slag. 

One class of wet methods aims to convert the zinc sulphide into 
zinc sulphite and sulphate, and to leach these with or without the 
addition of sulphuric acid. This is represented by Parnell,* Chen- 
hall, 5 Siemens and Halske, 6 Croselmire, 7 Lowe, 8 West, 9 and Le- 
trange. 10 Inthe Lower Hartz n (Prussia) the old process of heap- 
roasting the mixed sulphide ores and leaching out the zinc sulphate 
has lately received a fresh impetus from the manufacture of a white 
paint, called " Lithopon," which is a mixture of barium sulphate 
and zinc sulphide resulting from the double decomposition of 
barium sulphide and zinc sulphate. 

The other method of making zinc soluble is to convert it into 
chloride. Maxwell-Lyte 12 and Wilson 13 have based their processes 

1 Oesterreichische Zeitschrift fiir Berg- und Hiitten-Wesen, 1880, 
p. 462. 

2 Annales des mines, 1870, vol. xvii., p. 27. 

3 Engineering and Mining Journal, August 3, 1889. 

4 Berg- u. Hiittenmdnnische Zeitung, 1881, p. 254 ; 1883, p. 242. 

5 Ibid., 1884, p. 465 ; Iron, xxii., p. 465 ; Chemical News, xlii. 
p. 201. 

6 Wagner's Jahresberichte, 1888, p. 368. 

7 Engineering and Mining Journal, September 29, 1888 ; February 
9, 1889. 

8 Ibid., January 5, 1889, p. 17. 

9 Ibid., March 14, 1891. 

10 Berg- u. Hiittenmdnnische Zeitung, 1882, p. 489 ; 1883, p. 287. 

11 Zeitschrift. fiXr Berg-, Hutten-, und Salinen-Wesen in Preussen, xxv., 
p. 144; Berg- u. Huttenm. Zeitung, 1890, p. 131 ; 1891, p. 184. 

12 Engineering and Mining Journal, March 10, 1883. 

13 Chemiker Zeitung, 1884, p. 1670. 



SMELTING IN THE BLAST-FURNACE. 149 

on hydrochloric acid as a solvent. Electrochlorination has been 
tried by Slater. 1 

In smelting zinc-bearing by-products from refining works in a 
blast-furnace without ore the same slag is run over and over again 
until it becomes saturated with zinc. When a ferruginous slag 
contains 10 per cent, of zinc oxide it begins to be very mushy, and 
with 16 per cent, a gritty mass forms on top of the lead in the fur- 
nace, and the crucible cannot be kept open for any length of time. 
As far as the writer's experience goes, 16 per cent, is the maximum 
amount of zinc oxide that a singulo-silicate slag rich in iron can 
bear in a furnace having an Arents syphon-tap. 

k. Antimony. — Antimony occurs either as a sulphide or an 
oxide. The antimony sulphide behaves on the whole like lead 
sulphide, but is much more volatile. If decomposed by metallic 
iron, the resulting metal is more likely to combine with the lead 
than it is to form a speise with any excess of iron that may be 
present. It may also be volatilized. The oxide is generally pres- 
ent as an antimoniate of lead or iron, and this, being reduced to an 
antimonide, combines with the lead or the speise, if any is made, 
or with the matte. The two main injurious effects of antimony* 
therefore, are that it causes loss by volatilization and impairs the 
character of the lead. Antimonial speise is rare, and in making up 
an ore-charge no account need be taken of the small quantity of 
iron likely to be consumed by the antimony. 

Two difficulties have to be contended with in treating the anti- 
monial by-products of refining works in the blast-furnace. If the 
slag contains but little iron, thus requiring a high temperature, 
much antimony and lead are volatilized ; if rich in iron, some speise 
is liable to form, which either separates out, causing the loss of the 
antimony, or becomes mixed with the slag, making it rich. A fer- 
ruginous slag is generally preferred as the lesser evil of the two 
£ 119). 

1. Arsenic. — Arsenic occurs very frequently in argentiferous 
lead ores and must not be neglected in computing a charge. It 
causes loss by volatilization and combines with the lead, but not to 
such an extent as antimony, as it has a great affinity for iron and 
forms a speise. Arsenic combines with iron in various proportions, 
from IV. A- «ui. Speise having a composition thai lies between 
l> Ag and Pe 4 As is tine grained; it is not readily fusible, and not 
liquid when melted; it is likely to form hearth-accretions (" speise- 
1 Engineering and Mining Journal, June 2, July 14, 1888. 



150 METALLURGY OF LEAD. 

sows ") and to prevent a good separation of lead. Speise of the 
composition Fe 6 As has large cleavage planes, similar to spiegeleisen;. 
it is moderately fusible and pretty liquid when melted, does not 
readily form accretions, and retains only very few shots of lead. 
A speise of the composition Fe 6 As and upward seems to be just as 
harmful as one that contains less iron than Fe 5 As. In computing 
a charge, some metallurgists add anough iron to form Fe 4 As, keep- 
ing probably in mind that some arsenic is volatilized or combines 
with the lead; others assume Fe 5 As. In making up a charge with 
an ore that contains arsenic, this has to be considered in choosing 
the slag. By taking a readily fusible slag the speise, floating on 
the lead beneath the matte and far removed from the hottest part 
of the furnace, is liable to chill. This can only be prevented by 
selecting a slag that requires a great heat for its formation. Thus 
the fusibility of the slag must stand in an inverse ratio to the 
amount of speise formed. With large amounts, even a sesqui-sili- 
cate slag may be necessary. 

Sometimes old iron castings are added to loosen up any speise- 
sows that have been formed in the hearth. This is useful if the 
speise does not contain enough iron, as it supplies a deficiency, but 
does no good if the accretion is caused by too low a temperature* 

§ 55. Fuels Used in the Blast-Furnace.— The fuels used in 

lead smelting are coke, charcoal, and a mixture of the two. Experi- 
ments have been made to replace part of the coke by bituminous 
coal and anthracite; gaseous fuel has also been tried. 

a. Coke. — The coke to be suited for smelting purposes ought 
to be hard enough to bear the burden, slightly porous and low in 
ash. As regards the strength of the cell walls, Connelsville coke 
is the best, gas coke the worst. But Connelsville coke is very 
dense, and consequently does not oxidize readily. In order to over- 
come this, a high pressure of the blast is needed, which is on the 
whole not desirable. Gas coke might form a suitable fuel as regards 
its porosity, if the pores were not too large; they become partially 
clogged and the lumps of coke become glazed so as to be not only 
useless but harmful. A coke, therefore, that combines strength 
with a moderate degree of porosity is what is wanted. The dis- 
advantages of Connelsville coke are, however, more than made up 
for by the fact that it is not much broken up in transportation and 
by its low percentage in ash and its uniform quality. Thus, if coke 
has to be brought from a distance, Connelsville coke is the best. 

The amount of ash in coke varies from 10 to 20 per cent.; one 



SMELTING IN THE BLAST-FURNACE. 



151 



cubic foot of dry coke weighs about 50 pounds; it has about 50 per 
cent, of cell-space, and can bear about a 90-foot charge without 
crushing. 

COMPOSITION OF SOME COKES USED IN LEAD DISTRICTS. 





Connels- 
ville, Pa. 


Cardiff, 

Wales. 


Grand 
River, Col. 


El Moro, 
Col. 


Crested 
Butte, Col. 


Fixed carbon 


87.46 
0.49 

11.32 
0.69 
0.029 
0.011 

1. 


95.00 
0.01 
4.26 
0.68 
0.12 

l. 


93.75 

5.49 
0.76 
0.10 

1. 


87.47 

10.68* 
0.85 

1.85° 
2. 


92.03 
6.62* 

1.35° 
2. 


Moisture 


Ash 


Sulphur 


Phosphorus 


Volatile matter 

Reference 



Including moisture. * Generally higher. 
ANALYSES OF COKE ASH. 





Connellsville, Pa. 


El Moro, Ccl. 


South Park, Col. 


SiO, 


44 64 

25.12 

22.73 

6.95 

1.91 

2. 


84.50 
8.40 
7.10 

3. 


29.10 
23 10 
47.80 

3. 


Al a 3 

Fe 2 3 


CaO 


MgO 




■ 



1, M Mineral Resources of the United States," 1887, p. 396 ; 2, "Tenth Census of the 
United States," 1880, xv., p. 72 ; 3, Emmons, "Geology and Mining Industry of Leadville, 
Col.," p. 642. 

In computing a charge the ash of the coke has to be taken into 
account. 

Coke, used alone as fuel, gives clean slags and little fluedust. 

Before the coke is fed into the furnace all the fine pints 
lia\ c to be removed. This is done with a coke-fork having prongs 
h inch apart, while removing the coke from the sheds. Sometimes 
a scoop is used and the tines screened out by dumping the coke 
over a grizzlie, which discharges the coarse coke on the feed-floor 
and the tines into a bin, whence they are removed to be burned 
under the boiler. 

b. Charcoal. — As regards porosity, charcoal is the best fuel for 
a lead furnace, as it consists ' of a large cumber of small cells joined 
to each other by porous walls. Hence, being readily oxidized, it, 
is a good reducing agent for oxidized ores, and requires only alow 

1 Thorner, "Stahl nnd Eisen," vi., p. 71. 



152 METALLURGY OF LEAD. 

blast, which is an advantage. Its greater porosity is one of the 
reasons why weight for weight charcoal smelts more ore than coke. 
It also causes greater bulk (3 : 1), thus making the charge looser. 
This is favorable for quick smelting. The other reason is the low 
percentage of ash. The great disadvantage of charcoal is that very 
few kinds can bear any heavy burden; it breaks up and crumbles. 
Fine charcoal is not only worthless as a fuel, but it is a bad con- 
ductor of heat. It makes unclean slags and also causes loss in 
metal by increasing the amount of fluedust. 

Nut-pine (piflon) charcoal is the best, but it has to be well 
burned. Charcoal from lighter woods, such as yellow and white 
pine, quaking aspen, and cottonwood cannot be used alone in the 
furnace, and even with coke only a small percentage is allowable ; 
some metallurgists condemn it entirely. Mesquite makes a good 
charcoal, but it is obtained with difficulty in large pieces. Charcoal 
from hard woods, such as mahogany, cedar and oak, decrepitates 
in the furnace. When charcoal is exposed for any length of time 
to the open air it breaks up and the amount of fines becomes large. 
On the other hand, its quality is said to be improved by storing, 
through the oxygen that it absorbs. Lead smelters do not like to 
have large amounts of charcoal on hand. It should be stored 
where it is not exposed to the sun. 

Lightwood charcoal contains about 2 percent, of ash, consisting 
principally of alkali and alkali-earth carbonates, and does not affect 
a charge to any appreciable extent. One bushel weighs about 14 
pounds. The height of charge it can bear varies too much with 
the different kinds of charcoal to give a general figure; it is less 
than with coke. 

c. Coke and Charcoal. — From what has been said it is obvious 
that the ideal fuel for the lead smelter must combine the strength 
of coke and the porosity of charcoal; thus a mixture of coke and 
charcoal will put through more charges in a given time than either 
alone. The coke bears up the charge and prevents the charcoal 
from being crushed. This burns quickly, helps to ignite the coke, 
and, having hardly any ash, leaves hollow spaces for the blast to 
penetrate. 

d. Coke and Bituminous Coal. — Neil! 1 has succeeded in re- 
placing part of the coke by a non-caking or only slightly-caking 
bituminous coal, using separately lump, nut, and pea coal. He 
gives as results of his experiments that, besides the direct saving 

1 Trans. A. I. M. E., xx. 



SMELTING IN THE BLAST-FURNACE. 153 

in substituting the cheaper bituminous coal for the coke, jackets, 
slag and lead appeared hotter, the tuyeres brighter, and the cruci- 
ble kept open better. The slag assays ran lower than with the 
usual coke and charcoal mixture and the separation of slag and 
matte was good. On the feed-floor the charges settled more evenly, 
as fewer zinc accretions were formed, and the top was cooler than 
usual; while the volume of smoke was larger, there was no greater 
loss in the fluedust on account of the charge being cooler. In his 
furnace, 36 by 78 inches at the tuyeres, and 12 feet from the 
tuyeres to the charging-door, he uses coke and charcoal in the pro- 
portion of 3 : 1, and is able to have 27 per cent, of bituminous coal 
in his fuel charge. He expects to reach 33 per cent., and thinks 
that with a higher furnace one-half of his fuel might be bitumi- 
nous coal. 

e. Coke and Anthracite. — Dwight l publishes some experiments 
made by Rapp in substituting anthracite of goose-egg size for part 
of the coke, no charcoal being used. The furnace was 36 by 80 at 
the tuyere level and 9 feet from there to the feed-floor. The result 
was that the smelting power of the furnace was reduced as the 
proportion of anthracite was increased, e.g., one-third with anthra- 
cite as 60 per cent, of the fuel; otherwise the furnace remained in 
good condition. The top kept cool and the crucible open; there 
was a good reduction shown by a clean slag (0.4-0.8 $ lead), a 
matte low in lead (8 $ lead, 4 f copper), and a speise having a 
coarsely crystalline structure. Finally, less zinc accretions were 
formed than when coke alone was used, the charge containing 7.5 
per cent. zinc. 

f . Gaseous Fuel. — This has been used in a single instance. At 
the works of the Pennsylvania Lead Company 2 Blake introduced 
natural gas with the blast by inserting a gas-pipe through the 
tuyere-pipe. The amount of natural gas was regulated by stop- 
cocks and the blast-pressure increased so as to supply sufficient air 
for the combustion of the gas. Thirty per cent, of coke was suc- 
cessfully saved in this way. By replacing sixty per cent., the top 
of the furnace became too hot. That solid fuel cannot lie entirely 
replaced by gaseous is clear from the reactions going on in a blast- 
furnace that require solid carbon. For pecuniary reasons it is im- 
probable that any artificial gaseous fuel will ever be used in the 
blast-furnace. 

1 Trans. A. I. M. E., xx. 

* Trans. A. I. M. E., x\\, p. 661. 



154 METALLURGY OF LEAD. 

The weight of the fuel required in a lead furnace is generally 
expressed in terras of percentage of the total weight of the charge 
(ore, flux, and slag), sometimes by saying that one part of fuel 
bears so and so many times its weight of charge; the latter is more 
common in copper smelting. Misunderstanding occasionally arises 
from deducting the pounds of lead contained in the charge and re- 
ferring the percentage of fuel only to the slag and matte materi- 
al. Often the percentage of fuel used refers to the weight of the 
charge excluding the slag that is added. The amount of fuel re- 
quired varies with its character, with the fusibility of the charge, 
with the time of year, and the altitude at which the smelting is 
carried on. 

As regards the character of the fuel : coke that is rich in ash 
is not only an inferior fuel in proportion to the smaller amount of 
carbon it contains, but a considerable quantity of this carbon is 
consumed to melt the ash and the fluxes necessary to slag it. For 
this reason a smaller amount of charcoal than of coke would seem 
to be required. The exact opposite is, however, the case. This 
is probably because the charcoal crumbles and is crushed in its 
descent in the furnace. 

The richer an ore is in lead and the more fusible the rest of 
the charge, the less fuel will be needed. For instance, an ore con- 
taining zinc requires more fuel than one that is free from it ; a 
calcareous slag requires more fuel than one that is ferruginous ; a 
coarse and open charge requires less fuel than one that is fine and 
dense. 

In summer less fuel is generally required than in winter, not so 
much owing to the higher temperature as to the more rapid evapora- 
tion of the moisture contained in ore, flux, and fuel. The difference 
may be as much as 5 per cent. 

The altitude at which an ore is smelted makes a great difference 
in the amount of fuel required. Hahn, 1 for instance, gives the 
figures of 14 and 17 per cent, in Salt Lake City (4,000 feet above 
the level of the sea) as against 22 and 24 per cent, in Leadville 
(10,000 feet) ; the lower figure refers to the summer, the higher to 
the winter season. The only explanation that is at least in part 
satisfactory is the one given by Headden at a meeting of the 
Colorado Scientific Society, where the matter was informally 
discussed. A cubic foot of air entering the blast-furnace under a 
certain pressure will expand more at a high elevation, where the air 
i "Mineral Resources of the United States," 1882, p. 339. 



SMELTING IN THE BLAST-FURNACE. 155 

is rarefied, than at sea-level ; consequently more heat will be con 
sumed, and this has to be made up by an extra amount of fuel. In 
the same way more power and consequently more fuel is required 
at a high elevation to obtain this cubic foot of compressed air. 

Thus it will be seen that from various causes the percentage 
of fuel required in a blast-furnace must show great differences. 
For coke and a mixture of coke and charcoal (3:1 or 4 : 1) from 12 
to 16 per cent, is the common figure, rising sometimes to 18, rarely 
to 22. With charcoal alone it goes as high as 26 and 28 per cent. 1 

§ 56. The Roasting of Ores. 

a. Introductory Remarks. — The object of roasting ores pre- 
vious to smelting is to drive off as much sulphur and arsenic as pos- 
sible, thus reducing the amount of iron otherwise necessary to de- 
compose sulphides and arsenides and with it the amount of matte 
and speise, thus also permitting an increase in the amount of ore 
smelted. 

To decide whether it is necessary to roast an ore or whether it 
can be smelted raw, the character and amount of sulphides, the 
richness of the ore, and the cost of roasting have to be considered. 
As a general rule, any ore containing 15 per cent, sulphur is best 
roasted before it is smelted. To some extent the richness of the 
ore in silver may modify this rule. On account of the loss in 
silver endured -in roasting, it may in some cases be better to smelt 
an ore raw which contains more than 15 per cent, sulphur. An 
ore running 100 ounces silver to the ton is rarely roasted ; some 
metallurgists 2 draw the line at 50 ounces silver, which, however, 
seems rather low. 

The deleterious influence of blende in the blast-furnace has been 
previously emphasized (§ 54, j). If it is present to any extent, the 
ore will have to be roasted, for if it is smelted raw a large percent- 
age of slag would necessarily have to be added to the charge to di- 
minish the relative amount of blende and thus reduce its bad effect. 
In a general way it may be said that the higher the percentage of 
lead in a charge the more blende is permissible. For instance, if 
with mixed sulphide ores containing little or no pyrite the amount of 
lead present is twice that of zinc, the ore is smelted raw; if zinc and 
lead arc present in equal amounts, or if there is more zinc than Lead, 
it is best to roast the ore before smelting. 

1 Hahn, Oj). tit., p. 889. 

2 Newhouse, Engineering and Mining Journal, February 28, 1891. 



156 METALLURGY OF LEAD. 

TTith pyritic ores there is more margin. In smelting the ore raw 
in the blast-furnace pyrite consumes iron and thus reduces the 
furnace's capacity for ore; it forms much matte, consequently 
quite a percentage of lead and silver is not directly recovered in 
the form of base bullion. The matte has to be roasted before re- 
smelting. The iron previously consumed in the blast-furnace be- 
comes again available, so that the actual consumption of iron in 
smelting ores raw is not so great as is generally assumed. Whether 
a pyritic ore shall be roasted or not is decided by the percentage 
of sulphur it contains above that which is required for the amount 
of copper present. A large quantity of matte is, however, to be 
avoided, as the silver entering the slag increases with the percent- 
age of matte formed. This bad effect begins to show itself with 
10 per cent, of matte in the charge. 

Pure argentiferous galena ores rarely come in such quantities 
to a smelter as to make their separate treatment necessary. They 
are generally added to the charge in the raw state in such quanti- 
ties as to make up for the commou deficiency of lead. Impure 
galena ores are usually mixed with sulphuretted ores that are free 
from lead before roasting. It has already been shown (§ 9) that in 
roasting lead suljmide this is converted into oxide and sulphate, 
and that the presence of other sulphides increases the proportion of 
sulphate. Of the principal metallic sulphates 1 silver, iron, copper, 
zinc, nickel, cobalt, manganese, and lead sulphate, silver sulphate 
loses its sulphuric acid at a very low temperature, and lead sulphate 
gives it up only to a very small extent at a high temperature. If 
it is to be decomposed, the sulphuric acid has to be expelled from 
its combination by the stronger acid silica, which combines with 
the lead oxide and forms a silicate (§ 6 and 8). Zinc sulphate re- 
quires a bright -red heat and a considerable time to be converted 
into oxide. For this reason ores rich in blende are best roasted sepa- 
rately from those containing little of it. If several sulphides are 
roasted together, the order in which the sulphates give up their 
sulphuric acid differs from the one just stated, and will be iron, cop- 
per, silver, zinc, manganese, the silver being sulphatized again by 
the sulphuric acid set free through the decomposition of the iron 
and copper sulphate. 

The roasting of lead ores can be carried on in heaps, stalls, 
kilns, and reverberatory furnaces. So-called mixed ores consisting 
mainly of galena, pyrite, chalcopyrite, blende, and containing com- 

1 Kerl, " Grundriss der Metallhiittenkunde." Leipsic, 1881, p. 70. 



SMELTING IN THE BLAST-FURNACE. 157 

paratively little gangue, are sometimes roasted in heaps and stalls, 
the sulphurous acid being allowed to go to waste, or in kilns, when 
the sulphurous acid is to be converted into sulphuric acid. As this 
roasting is comparatively rare with lead ores, but very common 
with copper ores, and as the apparatus is practically the same, the 
method is best omitted here and the reader referred to the works 
on the metallurgy of copper 1 and the manufacture of sulphuric 
acid. 2 Here only the roasting in the reverberatory furnace will be 
considered. 

To roast an ore in the reverberatory it is charged at the cool flue 
end and gradually moved towards the hottest part, next to the fire- 
bridge, whence it is withdrawn as a pulverulent, agglomerated, 
or slagged mass, according to the prevailing temperature and the 
fusibility of the charge. As regards the subsequent smelting, it 
is best to slag the ore, as by obtaining the roasted ore in lump-form 
the disadvantages of treating fine ores in the blast-furnace would 
be overcome ; but other considerations often prevent this. The 
principal ones are the loss in lead and silver and the increase of 
cost. Newhouse 3 gives as a result of a series of experiments in 
roasting ores containing from 12 to 18 percent, lead a loss of from 
15 to 18 per cent, of lead and of 2 per cent, of silver with subse- 
quent fusion, and from 2 to 5 per cent, of lead and none in silver 
without it. By agglomerating the ore the loss will be only slightly 
higher than when it remains pulverulent. The sulphur will not 
be so effectually removed as when the ore is slagged, but more 
so than when it remains a powder. For instance, slag-roasted 
ore contains about 1 per cent, sulphur, while roasted pulverulent 
ore retains from 3 to 7 per cent, sulphur. The loss increases 
on the whole with percentage of lead in the charge. As a 
general rule, it may be said that an ore with 10 per cent, lead or 
can be safely slagged ; with from 10 to 20 per cent, it is ad- 
visable only to agglomerate it; should the lead run over 20 per cent, 
the temperature is besl kept so low that the roasted product remains 
pulverulent, or is only slightly adhesive when drawn from the 
furnace. 

Making up the sulphide ore-beds will therefore be regulated to 

1 Howe, "Copper Smelting," Bulletin No. 20, U. S. Geol. Survey, 
Washington, 1885 : Peters, " Modern American Methods of Copper Smelt- 
ing/' New York, 1891. 

8 Lunge, "Sulphuric Acid and Alkali/' London, 1891, vol. 1. 

8 Engineering and Mining Journal, February ^«, 1891. 



158 METALLURGY OF LEAD. 

some extent by the amount of coarse ore necessary to give the 
blast-furnace the required smelting power. Newhouse l advises 
the mixing of galena concentrates with pyrite, in order that the 
charge may contain 9 per cent, of lead, which is fused to obtain 
coarse lumps for the blast-furnace, and reducing the lead contents 
of the remainder down to 2.5 per cent., and roasting this without 
fusion, whereby sufficient lead for the blast-furnace would be ob- 
tained. At some works ores are fused that are entirely free from 
lead. In this case matte is liable to form in the fuse-box ; this, 
however, does not assist the volatilization of silver. 

The greater cost of fusing over roasting is caused by the extra 
labor, the larger quantity and better quality of fuel that is required, 
and the greater wear and tear of the furnace, owing to the high 
temperature and the corrosive action of slagged ore. 

All the statements made about the loss in lead and silver refer 
to the mixed sulphide ores treated by Western smelters, which as a 
rule run low in lead and high in silver. They are not intended for 
the pure galena concentrates free from silver (as in the Mississippi 
Valley) or low in silver (as in most European silver-lead works), 
because in both instances the ores are always slagged to a greater 
or less degree. The charges running 50 and 60 per cent, lead and 
free from impurities require when roasted a very slight increase 
of temperature to be slagged, care being taken to keep it as low as 
possible. Therefore the loss in lead and silver is slight, although 
the percentage of lead is high. Cramer von Clausbruch 2 states 
that at the Altenau smelting and refining works (Hartz Mountains) 
he obtains the best results in treating his galena ores if the charge 
contains 15 per cent, silica and from 55 to 60 per cent. lead. If 
there is more silica a base has to be added to effect a complete 
slagging at a reasonably low temperature; if there is less, some 
lead sulphate remains undecomposed. He notes the interesting 
fact that, if the roasted ore is not completely slagged but retains 
parts of sulphides and sulphates that have been only agglomerated, 
the silver and copper will be concentrated in the agglomerated 
part. One hundred parts of his charge give 85 per cent, thoroughly 
slagged ore, 10 per cent, of a mixture of slagged and agglomerated 
ore, and from 2 to 3 per cent, of unroasted agglomerated galena, 
the loss in roasting varying from 2 to 3 per cent. The slagged 

1 Loc. cit. 

8 Zeitschrift fur Berg-, Hutten- und Salinen-Wesen in Preussen, xxxi., 
p. 26 ; Engineering and Mining Journal, May 24, 1883. 



SMELTING IN THE BLAST-FURNACE. 159 

part of the charge contains one-half of the silver and only a trace 
of copper, while the other half of the silver and all the copper are 
concentrated in the rest of the charge. 

In making up charges for fusing-fu maces that contain 10 per 
cent, and less lead, quite different standards have to be followed. 
The principal base to combine with the silica of the sulphide ore 
will be iron, and the next important lead. Charges are made up 
so that they may be readily fusible and sufficiently acid not to 
corrode the bottom and side-walls of the fuse-box. The silica 
in them varies from 25 to 32 per cent., and the iron calculated as 
metallic iron is made to equal the silica. It is not common to add 
lime to a charge containing lead. With charges free from lead, 
proportions like 

Si0 2 FeO CaO 

32 32.. 24 

36 36 24 

are sometimes made up, resembling closely blast-furnace slags. 

The results obtained in roasting an ore depend not only on its 
chemical composition but also upon the size to which it has been 
crushed, the thickness of its bed in the furnace, the amount of rab- 
bling it receives, the time it remains in the furnace, and the temper- 
ature to which it is exposed. 

As galena oxidizes but slowly when heated with access of air, 
a large number of surfaces (fine-crushing) are necessary, if the 
roast is to have the desired result. Then, as the roasting proceeds 
in each particle from the surface to the centre, it is probable that, 
if the galena is too coarse, a reaction may take place where the 
oxide and sulphate found at the surface come in contact with un- 
decomposed sulphide at the centre, and the resulting metallic lead 
causes a considerable loss in lead and silver — another reason for 
fine-crashing. Ores that do not roast readily, i.e., ores rich in galena 
and blende, are crushed through an 8-mesh sieve. Ores that roast 
easily, •.//., pyritio ores and iron matte with 10 per cent, lead, are 
Crushed through a 4-mesh sieve. The oxidation with these is rapid, 
and the roasted product less fusible and more porous than it would 
be if richer in lead. 

The thickness of the charge on the hearth and the amount of 

wary working depend also upon the character of the ore. The 

richer it is in lead the thinner must be the charge and the more 

work does it require. The thickness of the bed varies from 3 to 6 



160 METALLURGY OF LEAD. 

inches, the rabbling being repeated from every J to every 1^ 
hour. 

The time required to roast the ore depends upon the readiness 
with which it is oxidized and its fusibility when roasted. Ores 
rich in blende require a considerable time and a high temperature 
before they are fused, if the zinc sulphide is to be completely con- 
verted into sulphate and this fully decomposed. Pyritic ores can 
be roasted quickly, and there is no danger of the half-roasted ore 
becoming sticky and adhering to the hearth of the furnace. Ores 
in which galena prevails require a very slow roast and a low tem- 
perature throughout, as even with the most careful roasting it is 
impossible to prevent the roasted ore from retaining undecomposed 
lead sulphide. 

b. Roasting Furnaces. — The furnace that has survived all other 
modifications of open reverberatories constructed for roasting is the 
" Fortschaufelungsofen " or open, long-bedded, calcining furnace. 
The characteristic of the American type is a roasting hearth from 
14 to 16 feet wide and from 40 to 60 feet long, with working-doors 
on either side. The hearth terminates at one end in a flue leading 
to the dust-chambers, at the other in a small vertical flue leading 
to the fuse-box or slagging hearth, which is about four-fifths as wide 
as the roasting hearth and one-sixth as long. The slagging-hearth 
with discharging doors on either side abuts on the fire-bridge. 

The old furnaces, about 6 feet wide, with working-doors only 
on one side, are probably not to be found now. Furnaces having 
two hearths, one on top of the other, can still occasionally be met 
with — for instance, at the large silver-lead works of Mechernich, 1 
Prussia. Having a double hearth has the advantage that the 
longitudinal extension is only one-half as great as with one hearth, 
and therefore, if cramped for space, one maybe justified in putting 
it up. But a double hearth requires a much more solid construction, 
and is therefore more expensive ; then if any repairing has to be 
done on the lower hearth, which is often the case, that part of the 
upper hearth situated above the place to be repaired has to be torn 
out to permit work. Finally, with a double hearth the workman, 
to turn over and move the ore on the upper hearth, has to stand on 
a high truck that runs on rails parallel with the furnace. Stand- 
ing on a shaky platform the man cannot do as good work as 
when he is on solid ground, and it is difficult to inspect his work. 

1 Berg- u. Huttenmdnnische Zeitung, 1875, p. 129 ; 1886, p. 434 ; En- 
gineering and Mining Journal, March 3, 1877. 



SMELTING IN THE BLAST-FURNACE. 161 

The consequence will be that all the ore is not moved towards the 
fire-bridge; particles will remain behind, and if the charge be 
rich in lead it will adhere to the hearth and gradually form a crust, 
which will have to be cut out. This requires shutting down the 
furnace in order to build a wood-lire on the hearth near the crust, 
in order that the flame may pass over it and soften it. It is claimed 
that fuel is saved by a double hearth, as much less heat is lost 
through the roof than is the case with a single hearth, but the 
same advantage can be gained by placing a layer of sand on the 
roof of a single-hearth reverberatory. 

It is a great improvement on the ordinary furnace to separate 
the roasting hearth from the slagging hearth (or fuse-box) ; this 
was first done in Colorado. It effects a sudden transition from 
powdery to really pasty ore, which is desirable. In the ordinary 
furnaces the hearth near the fire-bridge has a slight depression or 
sump, in which the roasted ore is melted down. The consequence 
of this is that the ore in front of the third and fourth doors from 
the fire-bridge is pasty, and if the fire has not been carefully 
watched, the heat may be excessive up to the fifth door. Not only 
does this interfere with a good roast, but it also renders the moving 
of the ore with the paddle a very arduous piece of work. To 
counteract this, it has been and often still is the custom with such 
furnaces to collect the ore from the third and fourth doors in a 
heap in front of the second door, and to melt it down into the 
sump, whence it is removed through the first door. 

This gradual passage from slagged ore through a sticky stage to 
pulverulent <mc is the reason why revolving roasting cylinders dis- 
charging directly into a stationary slagging hearth have, as a rule, 
been a failure. For instance, the " revolving cylindrical roasting 
furnace with slag hearth " of the old Swansea Silver Smelting 
and Refining Works, 1 described in every text-book, was a failure 
from the start. Bruckner 8 wanted to use at the German i a Works, 
Salt Lake City, 1'tah, two of his cylinders placed at right angles 
to each other. He proposed to set the upper one 4 feet higher 
than the lower one and to let it discharge into the latter by means 
of a chute. The lower one would discharge into a slag hearth, 
While the flame passed from the slag hearth through the lower and 
then through the upper cylinder. This plan was not carried OUt. 1 

1 Trans. A. I. M. /•:., iv., p. 42. 

* Engineering <m<i Mining Journal, January 15, 1887. 

» Terlmne, Trans. A. I. M. E., xvi.. v . 80. 



162 METALLURGY OF LEAD. 

Another combination of a Bruckner roasting cylinder and a station- 
ary slagging hearth is said to be at work at Spezia, 1 Italy. The 
cylinder, 15 feet long and 5 feet in diameter, is fired with producer 
gas and heated air. The gas-currents can be reversed as in a Sie- 
mens' regenerative furnace. By this means the ore is said to be 
uniformly roasted and sintering avoided. The roasted ore is dis- 
charged into a wagon and transported to the slagging hearth. The 
waste heat from this is used to warm the air for the roaster. 

c. Boasting Furnace icith Fuse-Box. — Figures 73 to 78 repre- 
sent the construction of the calcining furnace which is used at most 
Western silver-lead smelting works. It is generally spoken of as 
a 14 by 60 foot furnace, which refers to the dimensions of the 
roasting hearth. A detailed description is not necessary, as the 
drawings can be understood without it. A few remarks, however, 
may be in place. The roasting hearth is in four separate planes, 
divided by 3-inch offsets, which serve to keep the charges apart. 
The distance between roof and hearth is thus diminished by stages, 
leaving the former horizontal. This can also be done with a single 
inclined hearth, which is preferred by Hodges, 2 who gives it as his 
experience that the offsets, furnishing points of attack, lead to the 
injury of the hearth and are not required to separate one charge 
from another. Only that part of the roof above the lowest roast- 
ing hearth is built of fire-brick, the rest is of red brick. In the end 
view (Figure 76) are seen four openings for admitting air into the 
roasting hearth. The additional air required enters through the 
two doors next to the flue, which leads into the fuse-box, the door- 
lids being left slightly ajar. The working bottom of the fuse-box 
used to be (and is still sometimes) made of quartz sand, to which 
small amounts of slag are added after the sand has been put into 
the furnace and heated until it becomes slightly sintered on the 
surface. This bottom is represented in the drawing. It has not 
proved as satisfactory as was expected, and has been generally 
replaced by a 9-inch fire-brick bottom, built slightly concave. 
The bottom (Figure 78) rests on two arched roofs, and is thus 
cooled by air circulating below it. In the fire-bridge (Figure 
74) there is on one side of the air-space, a heavy cast-iron bridge- 
plate to bear the longitudinal stress of the hearth. The parts of 
the furnace that wear out fastest are the flue leading from the 
fuse-box to the roasting hearth and the fuse-box itself; the for- 

1 Engineering and Mining Journal, December 20, 1884. 

2 Engineering and Mining Journal, October 24, 1885. 



164 METALLURGY OF LEAD. 

mer is patched during the run with raw clay mixed with some 
burnt clay ; to repair the latter, the furnace has to be shut down. 
"Water-cooling of these parts has, as far as the writer is aware, 
not yet been tried. The cost of building a furnace, as shown in the 
drawings, is in Pueblo or Denver $3,000. The materials required 
are: 

Cast iron 12,000 pounds. 

Wrought iron 4,000 

Sheet iron 400 

Old rail buckstays 10,000 

Red brick 86,000 bricks. 

Fire-brick 15,000 " 

The tools required by each roaster-man are: 2 paddles (blade 
5 by 8 inches of \ -inch iron, handle 10 or 12 feet long of 1-inch 
iron); 2 rabbles (head 3 by 9 inches of -|-inch iron, handle 12 
feet long of f-inch iron); 1 slice-bar (1^-inch iron stem flattened to 
a chisel-point 3 or 4 inches wide), and 2 door hooks. The front- 
man has 2 scoops for the coal, 2 slice-bars, 3 rabbles (head 4 by 
9 inches, -^-inch iron; handle 10 feet, of -|-inch iron), and the 
necessary slag-pots to receive the slagged ore. 

In Figure 79 is given the plan of two furnaces forming part of 
a series of six. Near the flue end of the furnaces, 10 feet 8 inches 
above the furnace floor, runs a track, over which pass the ore-bug- 
gies to be discharged into the furnaces. At the opposite end of the 
building is another track at the same elevation for the coal-bug- 
gies delivering the necessary fuel into the bins. The flues of the 
six furnaces lead into a dust-chamber, 10 by 10 feet, up to the 
spring of arch. It is 500 feet long, and is connected through a flue 
4 by 10 feet with a circular brick stack w T hich has an inner diame- 
ter of 7 feet and is 85 feet high. The dust chamber has sliding 
doors to discharge the fluedust, two of which are shown in the 
drawing. 

d. Mode of Conducting the Process. — The mode of working a 
furnace with a fuse-box is pretty uniform at the different smelting 
works. The ore is sometimes dried on the roof of the furnace, but 
generally it is directly dropped, without drying, through the hop- 
per into the coolest part of the furnace, the weight of the charge 
varying from 2,400 to 3,300 pounds according to the thickness of bed 
the ore can bear. Sometimes the ore is shovelled onto the hearth 
through the last two doors, but this is only permissible when drop- 

















l ! 



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■ teSgfe 



rz 



fll 



s r M is * ' ill p _ *'i lii T— = ^ — ^-p — *'i isi ' is | 

r il r i r 1 r 




PLAN OF ROASTING-FURNACE HOUSE. 
B^ig. 79 






me 
"bin 
Wj 
not 
dra 
are 




5 b; 
iror. 
feet 
a cl 
mai 
9 ii 
nec< 

a se 
abo 
gies 
buil 



six 
spri 
4 b> 
ter 
doo] 
dra 1 
( 
fun 
wor 
gen. 
per 
van 
the 
thrc 









































3 i 






1 

















SMELTING IN THE BLAST-FURNACE. 1G5 

ping through a hopper is impracticable. The charge is spread uni- 
formly with paddle and rabble over the hearth at the highest point. 
If this is not separated by steps, but merely inclined, the charge is 
so spread that it shall lie in front of two doors. There it remains 
until the slagged ore is drawn from the fuse-box, when it is moved 
down the furnace to its second place on the next hearth, or in front 
of the next two doors. During, its journey to the fuse-box it is 
not only turned over with the paddle while being moved, but is 
raked with the rabble once, twice, or three times, according to the 
interval of time between the movings. Before the charge is trans- 
ferred to the fuse-box the latter receives some siliceous ore to 
protect the bottom. After dropping the charge the fire is urged. 
The liquefying begins at the surface, and much rabbling is required 
to bring the unf used parts from the bottom to the top. This is done 
at intervals of half an hour or fifteen minutes at a time at the be- 
ginning, later on for ten minutes, and towards the end for five. 
When the charge is fused it is drawn into slag-pots. This may 
be done in two ways, either all at once or in three instalments. 
The former method is used with charges containing 10 per cent, of 
lead, the latter with those running high, say 50 and 60 per cent, 
lead, for if they remain any length of time in contact with the acid 
hearth material they will take up silica and corrode the hearth 
unnecessarily. As soon as the front-man has drawn the slagged 
ore from the fuse-box the roaster-men drop in the next charge 
and begin to transfer the other charges downward, until the 
hearth near the flue is emptied and ready for a new charge. 
When the fused ore has become cold it is dumped from the 
slag-pots, broken up, and transferred to the feed floor of the blast- 
furnace. 

e. Results. — A furnace like the one shown in the drawings 
roasts and fuses in twenty-fours from to 6 to 8 charges (varying in 
weight from 2,400 to 3,300 pounds), consumes from 3 to 4 tons of 
bituminous coal, which is half lump and half pea, and requires 3 
nun in a 12-hour shift — one front-man, who attends to the fuse- 
box and firing, and 2 roaster-men, who do the work on the roaster 
hearth. If the ore is not fused two men are sufficient to do the 
work The cost of roasting and fusing with coal at $1.75 a ton, 
and labor, *2.2~) for 12 hours, is $2 a ton. 

f. Roasting Furnace without Fuse-Box. — The following is a 
description of a roasting furnace without a fuse-box where a 



166 METALLURGY OF LEAD. 

galena concentrate is slag roasted. It is at Mine La Motte, Mo. 1 
The hearth, 55 x 11-J feet on the inside, is slightly inclined from the 
flue to the bridge, making the respective distances from the hearth 
to the horizontal roof L5 and 22 inches. The top of the bridge- wall 
22 -J inches wide, is 9 nches above the hearth and 13 inches below 
the roof. The grate, 10 feet by 21 inches, is 3 feet 6 inches below 
the top of the bridge, this depth being necessary on account of 
the fuel used, which is wood. Of special interest is the construction 
of that part of the hearth where the ore is fused and of the bridge. 
The former is built into a wrought-iron pan resting on brick pillars 
between which the air circulates freely. It is formed by a full 
course of fire-brick, and is slightly concave. The air-cooling has 
proved very effective in preventing the corrosion of the brick. A 
similar result is produced by the air-flue passing through the bridge. 
When this was at first constructed in the usual way it was found that 
the middle part of the bridge was apt to be eaten through by the 
slagged ore. As a central wall divides the fireplace into two 
parts, the idea was conceived of closing the air-flue in the middle 
and erecting a small chimney on the roof, communicating with 
both parts of the air-flue. By this means a strong current of air 
could be passed through the flue. The experiment was a success, 
and the improvement has now been used for ten years. 

The writer, has used a water jacket in fusing antimoniate of 
lead in a reverberatory furnace, and has found that it stopped all 
leakage at the bridge. 

The furnace is charged every 6 hours with 2 tons of galena con- 
centrates, to which some sand is added as acid flux. The thickness 
of bed is 6 inches. The galena runs from 40 to 70 per cent, lead, 
and from nothing to 25 per cent, iron, and is crushed to pass a 12- 
mesh sieve. The slagged ore retains from 4 to 6 per cent, sulphur ; 
4 men work on a 12-hour shift, and 0.42 cords of wood are burned 
per ton of ore ; a second roasting furnace, without fuse-box, is de- 
scribed and illustrated in § 76.4. 

g. Products. — The two products obtained by roasting are 
roasted ore and fluedust. The composition of some European ores 
running high in lead that have been agglomerated or completely 
slagged is given below. Complete analyses of slag-roasted ores 
running about 10 per cent, lead are not made at the smelting works, 
and therefore cannot be quoted. 

1 Private communication from J. T. Monell, May, 1891. 



SMELTING IN THE BLAST-FURNACE. 



167 





Rodna 


. Tran- 


Mechernich, 


Freiberg, 


Hall Val- 


Mine La 




sylvania. 


Prdssia. 


Saxony. 


ley, Col. 


Motte, Mo. 


Raw Ore. 


Ag-glomer- 
ated Ore. 


Raw Ore. 


Slagr- roast- 
ed Ore. 


Slag- roast- 
ed Ore. 


Roasted 
Ore. 


Agglomer- 
ated Ore. 


Pb 


47.29 


54.27 


60.40 


62.08 






75.95 


PbO .... 








.... 


22.6 


42.04 




Ag 


0.059 


6.061 


0.0105 


.... 


0.13 


Ag o O,0.21 




Au 


0.0001 


0.0001 


.... 






.... 


Ni-Coi.ll 


Cu 


trace 


0.02 


0.17 


6.14 








CuoO.... 




.... 




.... 


6.3 






CuO .... 




.... 




.... 




i.7i 




As 


0.34 


0.030 


Ni6,10 


n. d. 








As 2 5 .. 










1.1 


.... 




Sb 


6.02 


6.027 


6.07 


6.08 








Fe 


20.36 


24.06 


0.80 


0.56 


.... 






FeO . . . . 




.... 








3.59 


4.5 


Fe,O s . . 










33.3 






Zn 


0.67 


0.87 


6.15 


n. d. 


16.0 






ZnO .... 


«... 


.... 






.... 


.... 




A1 2 3 .. 


0.11 


0.23 


3.60 


4.24 


1.8 


8.11 




CaO .... 


trace 


trace 


0.88 


1.28 


2.0 


0.42 


0.76 


MgO . . . 


trace 


trace 






0.5 






BaO 










.... 


12.05 




SiOo.... 


0.49 


6.80 


22.05 


22.77 


17.4 


22.71 


7.21 


co 2 . . . . 


.... 




.... 


.... 






.... 


so 3 




2.25 






trace 


6.31 




s 


29.86 


2.72 


9.72 


6.60 


3.6 


2.94 


7.41 







13.41 








.... 




Reference 1 


I. 


1. 


2. 


2. 


3. 


i. 


5. 



References : 1, Oesterreichisches Jahrbuch, xxix., p. 27 
tung, 1875, p. 129 ; 3, Oesterreichisches Jahrbuch, xvi., p. 397 ; 
School of Mines Quarterly, ix., p. 216. 



2, Berg- u. 
4, Trans. ± 



Hiittenmcinnische Zei- 
. I. M. E., v., p. 568 ; 5, 



The amount of liuedust carried off with the gases is about two per 
cent., and is practically all collected in the dust chambers; of the 
metal volatilized in fusing very little if any is recovered. Fluedust 
from furnaces when the ore is simply roasted has a brownish color ; 
if slagging is carried on, it is gray from volatilized lead and zinc. 
The composition of liuedust from some European works is subjoined. 



Pb 

Ae;, oz. per 

< 'n 

Od 

As 



ton. 



Zn.. .. 

Fe.O, 
OaO... 

MgO . 

SiO., . 

so 8 .. 

s .... 
H 2 0.. 



Friedrichs- 1 
hiitte, Silesia. 


Freiberg, 1 Saxony. 


58.0 


26.27 


21.37 


ni.-j; 


2:5.33 


.... 





6.6a 


.... 


7.56 


87.' 5 


46.41 


2.5(5 


19.10 


2.11 


0.45 






4.86 


1.68 


1.96 


0.80 


0.63 


0.48 




0.60 


0.35 


0.15 




8.90 


4. si 


3.58 


25. HO 


2S.14 


8.28 


7.05 






1.84 


0.60 






8.40 


3 ;-> 


1.60 




i .46 


8.37 



Herlng, " Die Wrdic-htun;,' iles lliit icnruiidi.-" Stuttgart, 1888, p. 34. 



168 METALLURGY OF LEAD. 

lies x gives as the average assay of fluedust obtained from 
roasting and from fusing ores the following figures: 





ROASTING. 


FUSING. 


Ag 

An 

Pb 

Zn 


23-28 
0.5-1.5 

5-10 
none 


20-37 

0.10-0.75 

19-40 

7-8 



The large quantity of gold contained in it is probably due to 
gold-bearing pyrite concentrates from Colorado gold mill, used as 
iron flux. The manner of working fluedust is discussed under the 
head of fluedust from blast-furnaces, § 85. 

§ 57. The Selection of a Furnace Site. 2 — In selecting a 
furnace site, economical as well as technical considerations must 
come into play. Some of these are: 

a. Are the conditions such as to justify the erection of a per- 
manent plant, or is the structure to be only a temporary one ? If 
the former, the plant is built of stone, brick, and iron; all arrange- 
ments are so planned as to allow enlargement by simply adding to 
the old j)lant. If the plant be a temporary one, as little money as 
possible will be spent on the building, which is usually a light 
wooden one. Often the necessary capital is not available to build 
properly from the start, and a temporary structure has to grow by 
its own profits into a large, permanent plant. In such a case the 
general plan will commonly be the same as that of the per- 
manent plant, the temporary structure being replaced as opportu- 
nity permits, although in some instances it may prove more advanta- 
geous to erect the permanent plant independently of the temporary 
one, this continuing to run till the other is ready to be started up. 

b. What is the precise nature of the work to be done ? Do the 
ores arrive in large or small quantities, at regular or irregular in- 
tervals, and through the whole year or only during a part of it? 
Upon these points it will depend whether the ore, as it comes, can 
be directly conveyed to ore-beds, whether bins are sufficient to store 
it, or whether a storage yard has to be provided. If the last, the 
ore may generally lie uncovered, but sometimes it has to be pro- 
tected from snow. For fuel, protection is always needed. The 
size of the plant and the richness of the ore will determine whether 

1 Engineering and Mining Journal, January 30, 1886. 

2 Locke, J. M., " Smelting Plants," Cincinnati, 1883. 



SMELTING IN THE BLAST-FURNACE. 169 

hand sampling or mechanical sampling or a combination of the two 
is most desirable. The character of the ore must decide whether 
it will be necessary to provide a place for ore-roasters and their 
crushing plant. 

c. What is the character of the ground on which the works are 
to be erected ? A hillside is always chosen, if possible. The 
questions to be settled are: how much fall is desirable and how 
much can be had? With an insufficient fall the main stress will 
be laid on having at least two terraces for the ore floor and the 
furnace floor with a convenient slag dump. In some instances 
works have to be built where there is no fall whatever, which re- 
quires additional elevators. 

If excavating has to be done, the nature of the soil must be 
considered. If it be rocky, it requires blasting ; if earthy, grading 
with a scraper will be sufficient, but will necessitate the erection 
of retaining walls. 

d. What water suj^ply can be had for boiler and jackets? If 
there is no natural flow from an elevated point, a well will have to 
be sunk or the water obtained from a creek below, which requires 
pumps and additional boiler power to raise the water to the tank. 
The character of the water, if hard or soft, clear or muddy, may 
also be an important consideration. 

§ 58. General Arrangement of a Smelting Plant. 1 — In lay- 
ing out a smelting plant, the principal aim must be to simplify as 
much as possible the handling of materials. This is done in two 
ways: first, by utilizing every possible opportunity to discharge the 
materials by gravity into and from the truck, when transported from 
one place to another ; second, by making the run-ways as short as 
possible without too much crowding of apparatus, which must be 
avoided on the all-important score of necessary ventilation, and 
also to give room for moving freely about. 

If there is ;i good natural fall, the arrangement will be as fol- 
lows: taking the furnace floor as base-line, there will be on one 
side the slag dump, with a fall of 20 or more feet; on the other, 
the furnace, reaching to the feed floor, 16 or more feet above. The 
roasting furnaces, will be on the third level, if the ore is all to be 
roasted; otherwise at the height of the l'eed floor. Next comes the 
track from which the crushed ore is discharged into the hoppers of 
the roasters, 8 or more led above the roaster floor and below tin; 

1 Locke, J. M., "Smelting Plants," Cincinnati, 1883. 



170 



METALLURGY OF LEAD. 



discharge of the sampling mill, through which most of the ores 
that come to the smelter pass. It will be the level of the ore yard 
and the storage place for fluxes and fuels. 

The precise way in which the floors are placed will vary with 
the configuration of the ground, as it is necessary not only to have 
the right fall, but also convenient grades for bringing in the mate- 
rials and carrying away the products. 

The situation of the machinery for driving the blowers, eleva- 
tors, and dynamos, and that of the pumps and of the machine and 
blacksmith shops, also varies, although it is usually on the furnace 
floor. 

The machinery for crushing is all found in the sampling de- 

Kig.. so 



■-■'■....■irmn,,,. 


" 1 l"" rni 


^r """:^ l7 "■"""■' 






uuu4iix — -" " 


""-..I..T*. 


-: : ;"|" 






R. R. 












partment. The steam is usually all furnished from one set of 
boilers. 

The office and laboratory are ordinarily near the ore yard. 

The general arrangement of the works of the Globe Smelting 
and Refining Company, at Denver, Col., as planned by the super- 
intendent, Dr. M. W. lies, is based on the following scheme: if 
two intersecting lines are taken, running east and west and north and 
south, the ground-level will be represented by the southeast and 
southwest fields, containing respectively the furnaces facing south 
in a row, and the boilers and machinery for blowing, lighting, etc. 
The upper fields will show in the northeast field the calcining fur- 
naces, and in the northwest field the sampling department, on the 
level of the feed floor, the only other level. Two sets of tracks on 



SMELTING IN THE BLAST-FURNACE. 



171 



the upper level, running east and west, bring in ore, flux, and fuel 
on either side of the sampling and calcining departments; another 
track on the furnace floor takes away the base bullion produced. 
An inclined elevator running north and south brings the foul slag, 
matte, and fluedust from the furnace floor to an elevated track be- 
tween the calcining and sampling departments and dumps the 
three products in the places where they are to be further treated, 
i.e., the slag near the ore beds, the matte near the sampling mill 
to be crushed before roasting, and the fluedust near the fusing 
furnaces, where it is to be slagged. 

Figures 1 80 and 81 represent the general arrangement of the 
smelting department of the Grant and Omaha Smelting and Refining 

Fig. 81 




>»^^K^^^^^^^^^^B 



Company's works at Denver, Col. A few changes in detail have 
bees made, but the general outline is correct. This department 
corresponds to that one of the Globe Works represented in the 
southeast field just referred to, and is similar iii construction. 

LEGEND. 

Nog, 1-8 — .Blast-furnaces. 

a = Projection in the dust chamber (abolished). 

r = Bustle-pipe. 

s = Induction-pipe. 

v = Flue carrying fluedust from the blast-furnace into the dust 
chamber. 

ir = Sh«'ft-iron curtain through which the charges are fed into the 
blast-furnace (abolished). 

1 Frost, Engineering and Mining Journal, March 24, 1883. 



172 METALLURGY OF LEAD. 

y = Elevated tramway for the fuel trucks that the contents may be 
dumped on the feed floor. 

z = Telescope stack, used in blowing in or out, to carry explosive or 
hot gases out into the open air (abolished). 

RR = Broad-gauge track on feed floor on which ore and fluxes arrive. 

TR — Narrow-gauge track on furnace floor delivering the base bullion 
to the broad-gauge track R. R. 

TR' = Tramway for raising slag, etc. (altered). 

To be noted especially is the large ore-bedding floor. It is the 
practice to make two large ore beds, each occupying nearly one- 
half of the floor, about 8 feet high and holding about 3,000 
tons of ore. All the furnaces receive their ore from one bed, and 
while this is being consumed, the other bed is made. 

Another feature not to be overlooked is the position of the 
fronts of the furnaces in regard to the points of the compass. 
Facing northwest they are as much as possible in the shade, an im- 
portant consideration in hot weather. 

The sheet-iron hood z', placed in front of each furnace to carry 
off the fumes that arise on tapping the slag, never did its work 
satisfactorily. The hood now ends in a horizontal pipe (x\ Figure 
96), which either terminates in the dust-chamber or in a galvanized 
iron pipe, common to a number of furnaces, and connected with a 
fan which sucks off the fumes and discharges them into the open 
air. Sheet-iron plates hung on either side of the hood prevent the 
draught on the furnace floor from carrying off the fumes into the 
building before they can be taken away by the hood. 

The works of the Montana Smelting Company, at Great Falls, 
Montana, planned by Mr. A. Eilers, are shown in plan and vertical 
section in Figures 82 and 83. Their general arrangement differs very 
much from that of the two works previously discussed. In the latter 
the longest extension is parallel with the row of furnaces ; at the 
Montana works this is reversed. A striking feature is the extensive 
roasting plant with its separate ore yard. The study of the plans 
will show that great attention has been given to the handling and 
storing of large amounts of ore. The loaded cars, arriving on the 
main track, are Aveighed on track-scales and then switched off on 
side tracks, which lead to the " crushing house and sampling 
works," to the "bins for sulphide ores" (tracks 9 and 10), and 
to the "bins for carbonate ores" (tracks 2-6). At the sampling 
works all ores requiring crushing are received, and the sulphide 
bins are filled with concentrates, which are sampled by fractional 




:t 



a £ ^ S a 

f f I f § 



» J1 

! i 



a H I dOHS 

3 JJ (iHiiwsxovna f | 
» ■' : ' 



J 




t - on xovai 
'i'oVxovax" 



I SOL— 

$'a~sno}{ tuipuvog oj/—* 



ol 



,o>** 



Mfttg 7.>DJJ«_ 



CT- 



Hyot 



3--E 



'"h -on *ovSj!" 
(U'QNJSOKblX 



— 63HO aaiHenos «oj snib 



•A« N«3Hi.UON *10 



onv asnoH ONiHsnbo 

xdvMV3snoH w 9Nndwv8 



174 METALLURGY OF LEAD. 

selection while they are being unloaded. The lower bins are for 
ores, sampled in the same way, that go straight to the blast-furnace. 
Track No. 1 brings the coal for the boilers and takes the base 
bullion produced in the blast-furnaces. 

The crushing house and sampling mill, which is fitted up with 
the necessary machinery for crushing, sampling, and grinding, de- 
livers the sulphide ores over a tramway to their respective bins. 

In the roaster building is room for twenty calcining furnaces 
(see § 76.4) and fusing furnaces. They deliver their gases into two 
parallel flues (see Figure 163) running along the centre of the build- 
ing (Figure 82). These are built one against the other and are com- 
bined into one main flue after they have received the gases of the 
single furnaces. This passes out of the building and terminates in 
a series of dust chambers, like those in Figure 171, connected with 
the stack by a small flue. The tramways are for carrying ore, fuel, 
and ashes. The floor of the roaster building is on the same level 
as the top of the lower ore bins; the roasted ore can thus be easily 
discharged into them. 

The blast-furnace building shows three floors instead of two, as 
is usual. The upper or ore-bedding floor is on the same level as 
the bottom of the bins for carbonate ores. From the beds made 
on it the ore mixture is dumped onto the feed floor near the single 
furnaces, where the charges are made up. The fluxes, arriving on 
track No. 2, are passed through the crusher, if necessary; the fuels 
arrive on the same track. On the furnace floor is room for five 
blast-furnaces, the details of which are shown in Figures 99-120. 
The furnaces face toward the river, where the ground is high and 
furnishes an ample slag dumj). The furnace gases pass through a 
wide flue back of the furnaces, leading into a series of dust cham- 
bers, which are connected with the stack. The details of flue and 
chambers are given in Figures 171-177. The ventilation of the 
furnace floor shown in Figure 83 is the one strongly recommended 
by Eilers. It consists in allowing the feed floor to extend only to 
the front of the furnaces and closing it off from the entire front 
part of the building by a wooden partition, which slants backward 
to the ridge-pole of the roof. Thus a hood as long as the furnace 
building, and having a width of about 12 feet, reaches from the 
level of the feed floor to the ventilator and draws off all the vapors 
and smoke from the tap-hole, the slag-pot, and the lead-well. A 
similar arrangement at the smelting-works in Clausthal, 1 Prussia, 
1 Private notes, 1890. 



SMELTING IN THE BLAST-FURNACE. 



175 



was very efficacious in removing the danger of lead poisoning, which 
had previously been a common thing. 

Next to the furnace room are placed the boilers, engines, blow- 
ers, etc. The drawing, Figure 82, shows the blacksmith-shop, the 
pump-house, the water pipes and fire-plugs, the water-tank and 
lower reservoir, which receives the overflow of the jackets, but 
not the upper reservoir. 

§ 59. The Blast Furnace and Its Accessory Apparatus. 1 

a. Introductory Remarks. The blast-furnaces used in lead smelt- 
ing are of the most various description. Taken in cross section 
they are square, polygonal, circular, oblong, and elliptical. In 
vertical section they are prismatic or with sides tapering towards 
the bottom, and in addition may or may not have a bosh. Then, 
the smelting zone may be enclosed by water jackets and the crucible 
may be internal, or partly internal and partly external. Furnaces 
with a detached crucible have been constructed and patented, 2 but 



BLAST FURNACE 
BUILDING 




REVERBERATORY FURNACE BUILDING. 



r ^ Lr- j " ! IT 



rpr 



f-io- 



:FiK. 83 



CRUSHING HOUSE AND 
SAMPLING WORKS. 



J Lll 1. «r-l. .1 1 

159-6- 



are not, as far as the writer is aware, in use. Finally, the lead may 
either be tapped from the bottom of the crucible or removed by 
means of Arents' automatic tap. 

Only two kinds of furnaces are now in use in this country, both 
having an internal crucible, water-jackets and Arents' automatic 
tap. The one is circular, has the form of an inverted truncated 
cone, and is used for smelting small quantities of ore or by-products 
of refining works ; the other (Figures 84-120) is oblong, its sides 
are either vertical or only slightly inclined, it always has a bosh, 
and is the common ore-smelting furnace. Oblong furnaces have 
almost entirely replaced circular furnaces. 

In ihr subjoined Legend the Letters used refer to the similar 

1 Working tools are discussed in §68-67. 

•Devereux: L887, December 6, No. 374,239; 1888, April 17, Nos. 
881,118 and 881,119; June 12, No. 884,849; 1889, July 23, Nos. 407,886, 
407,886, and L07.837; December 17, Nos. 417,814 and 417,810 ; 1890, May 6, 
No. 127,058. Konemann : 1888, October 9, No. 390,785. Wilson : 1889, May 
21, No. 408.815, and Qthers. 



1TG 



METALLURGY OF LEAD. 



parts of the four blast-furnaces (Figures 84-120) chosen as charac- 
teristic types. 



LEGEND. 



Cast-iron breast jacket, right. 

Cast-iron breast jacket, left. 
Shaft, red brick shell. 

Shaft, fire-brick lining. 
Cast-iron side jacket. 

Wrought-iron side jacket. 
Wrought-iron back jacket. 
. Wrought-iron front jacket. 
Slag-spout. 
Crucible. 
Lead-well or basin of Arents' 

automatic tap. 
Syphon, the inclined channel of 

Arents' automatic tap. 
Water jackets. 
Cast-iron water feeder. 
Lateral water-supply pipe. 
Water- feed pipes. 
Water-overflow pipes. 
Galvanized- iron water trough. 
Cast-iron drain pipe. 
Breast of furnace. 
. Tap-hole. 
Tapping jacket. 
Tuyere. 
Blast-pipe. 
Wind-bag. 
Bustle-pipe. 
Induction-pipe. 
Cast-iron collar, supporting — or 

carrier-plate. 
Cast-iron pillar. 
Down-comer. 

Sheet-iron curtain. 
Main water-supply pipe. 
Lugs fastened by bolts. 
Sheet-iron hood. 

Lead-spout. 

Wrought iron bolts. 
Corner-irons for tie-rods. 

Tie-rods. 



Interior of water jackets. 
Channel of water feeder. 
, Brick arch. 
, Crucible-castings. 

Strengthening-ribs of crucible- 
castings. 

Telescope stack. 

Chains. 

Counter-weights. 
. Feed-door. 

I-beams. 

Capital of pillar. 

Brass nozzle. 

Hand-hole. 

Wrought-iron pipe connecting 
the single jackets. 

Eye or peep-hole. 

Cast-iron flange. 
, Hog-chain. 

Iron band. 
. Top-plate. 

Pipe leading to fan. 

Bed-plate. 

Hood leading into pipe x'. 
. Tuyere-box. 
. Steel rails. 
, Wrought-iron rods. 
. Expansion-space. 
. Sliding sheet-iron door. 

Angle-iron ring. 
. Angle-iron damper. 
. Crank with nut. 

Circular guide with slot. 

Groove for damper. 
. Water feeder. 

Peep and poking-hole. 
'. Outlet for furnace gases. 
. Swinging valve. 
. Cross-pin. 
. Movable weight. 



Figures 84-88 represent the blast-furnace of the Omaha and 
Grant Smelting and Refining Company's Works at Denver, Col., of 



SMELTING IN THE BLAST-FURNACE. 






1883 ; Figures 89-95, the furnace of the Globe Smelting and Re- 
fining Co., at Denver, Col., 1891, designed by lies ; Figures 96-98, 



Fig.„84 



Trig. 85 




a furnace designed by the Colorado Iron Works at Denver, L891 ; 
Figures 99-120, the furnace of (he Montana Smelting Co., Great 
Falls, Mont., 1891, designed by Eilers. 



178 



METALLURGY OF LEAD. 



No drawing of the circular furnace is given, as it is little used 
now ; its advantages and disadvantages will, however, be dis- 
cussed in connection with the oblong furnaces. 

The materials required for the erection of a furnace, as shown 



Ficr. so 




SIDE ELEVATION. 
BLAST-FURNACE OF THE GLOBE SMELTING CO.'S WORKS, DENVER, COL. 

in Figures 89-95, are : cast iron, 27,300 pounds ; wrought iron, 3,200 
pounds; steel beams, 4,250 pounds ; fire-brick, 9,500 bricks ; red 
brick, 17,000 bricks. Where a "telescope" stack is used, 1,600 
pounds of wrought iron must be added to the above figures. The 



SMELTING IN THE BLAST-FURNACE. 



179 



cost of erecting the furnace, excluding all the fittings for blast and 

water, at Denver, was $1,200, one-quarter of which went for labor. 

b. foundation. — The first thing in erecting a furnace is to 

have a solid foundation. Its depth will depend on the character 



Fig. 90 






FEED FLOOR 



JOIST 




SECTION. 

BLAST-FURNACE OF THE GLOBE SMELTING CO.'S WORKS, DENVER, COL. 

of the subjacent -round. If there is exposed bed rock this will 
furnish as good a foundation as can be wished for. If there is 
loose soil or gravel covering bed rock for not more than L0 feet, it 
is best to excavate until this is reached; otherwise a depth of 5 
feet will usually be sutlicient to start the masonry below frost line, 



ISO METALLURGY OF LEAD. 

and to give the foundation the requisite strength. With very- 
loose soil it is sometimes advisable to place in the bottom of the 
pit two layers of 3- or 4-inch planks spiked crosswise to each other, 
and upon that to build the foundation, which should extend from 
2 to 3 feet beyond the bed plate and the four pillars. It is built 
up of undressed rock, well rammed into place, the largest pieces 
being used for the corners, and care being taken to till up the 
crevices and joints with as many spalls as possible ; the whole is 
well grouted with a mixture of four parts of lime mortar and one 
part of cement. The topmost course must be absolutely smooth 
and horizontal, being generally of brick. 

If one furnace is already in operation and a second one is to be 
erected, the simplest way of obtaining a good foundation is to 
empty the liquid slag into the place that has been excavated, or to 
throw in broken up slag and to cement it with liquid slag. The 
top is evened and levelled by making shallow rectangular areas 
surrounded by sand or pieces of iron rails, and filling them level 
with liquid slag. Any ridges or other rough parts that remain are 
removed by chipping. 

On the foundation is spread a thin clay mortar, upon which the 
wrought-iron bed plate y[ (Figures 89, 90, 96, 98, 99, 100) is 
placed. 

c. Shaft. — On the foundation are erected the four hollow cast- 
iron pillars v. (Figures 84-86, 89, 90, 92-94, 96-101), which are to 
support the shaft. 

The height of the shaft, i.e., the distance from the centre of the 
tuyeres to the feed floor, has been somewhat increased of late. It 
used to be from 10 to 12 feet; it was then increased to 14 feet, 
which is the common dimension now, although occasionally it 
reaches 18 feet. The increase of height has been necessitated by 
the greater pressure of blast required for the highly siliceous and 
calcareous slags. The ferruginous slags, formerly made, needed 
only a pressure of from J to 1 inch quicksilver, instead of from 1 
to 2 inches as at present. By enlarging the distance between the 
tuyeres, to increase the capacity of the furnace, it became necessary 
also to increase the pressure from 2-2^ inches quicksilver. This is 
a disadvantage, as will be shown further on. 

The horizoyital section of the shaft is either a circle or an ob- 
long. When a square or polygonal furnace is blown out, the inside 
will have a similar appearance to that of the circular, as the corners 
very soon fill up. The same holds good with the oblong furnace. 




;%}- V U lbolt U V 



u^^i^A 




iy 2 hole J" 



PLAN THROUGH WATER JACKETS. 



Fig. 94- 

DETAIL OF COLUMN. 







og n og 



W 



8° 81 



q 



an 



SMELTING IN THE BLAST-FURNACE. 181 

The circular furnace gives, as regards the quality of work, sastisfac- 
tory results; there is an even distribution of blast and heat, and as 
it offers the largest surface for the smallest circumference, the loss 
of heat by radiation is the least possible. The drawback lies in 
the quantity of the work, which is limited, since the diameter 
at the tuyere-section must not exceed 42 inches (36 inches being 
the common dimension), as too great a pressure of blast would be 
required. For large quantities of ore the oblong form is therefore 
the proper one, as the area can be enlarged by making the furnace 
longer without increasing the distance between the tuyeres. Thus 
the length of some oblong furnaces has been doubled in the last 
ten years (from 60 to 120 inches). In increasing the width the 
pressure of the blast has to grow, and with it the height of the 
furnace. "With high -pressure blast the heat creeps up and smelting 
begins at the top of the jackets, in which case the walls are liable 
to be eaten out, instead of just above the region of the tuyeres, 
which of course becomes cool and causes rich slags and rich mattes. 
Another result is a greater loss of metal, as the retarding effect of 
the boshes on the hot gases becomes weakened and the working 
height of the furnace diminished. Furnaces have been built 48 
inches wide at the tuyeres, the water-cooled nozzles protruding 6 
inches into the furnace, thus making the distance between the 
jackets 60 inches. In order to prevent a wedge of unmelted charge 
from forming along the centre, the bulk of the fuel has to be fed 
there in order that the powerful blast, which has more than 2$ 
inches quicksilver pressure, may penetrate the charge. If a furnace 
60 inches wide at the tuyere-section smelted twice as much charge 
as do two 30-inch furnaces, its size might be justifiable, as in the 
end it might be cheaper work. But it does nothing of the kind. 
It puts more charges through than one set of furnace hands can 
handle in a twelve-hour shift, and not enough for two sets, and is 
more expensive than smaller-sized furnaces. It is improbable that 
60-inch lead furnaces will again be built. 

For the usual run of ores the width at the tuyeres may range 
from ;;<> to :;<; inches, the height above the tuyeres from 11 to 14 
feet, and the blast pressure from 1 to l-£ inches quicksilver. A 
furnace 33 by I00 inches at the tuyeres, with five :\\ -inch tuyeres on 
either side, 12 feel active height, will smelt with l-j- inches quick- 
silver pressure as much medium-coarse charge as one regular crew 
of men will be able to handle in a twelve-hour shift, i.e., 60 tons 
of charge, or about 45 tons of ore. 



/ 

182 METALLURGY OF LEAD. 

The vertical section of all oblong furnaces shows the bosh, and 
with it the contracted tuyere-section. This last secures a more 
perfect and rapid combustion and thus a concentrated, intensified 
heat, with the result of a quicker fusion and a more complete de- 
composition of sulphide and arsenide of lead. If somewhat higher 
up the width of the furnace is suddenly enlarged by the bosh, the 
zone of fusion will be narrowed; further, the gases generated at the 
tuyeres will be evenly diffused, thus checking the velocity of their 
upward motion; by gradually giving up their heat they prepare the 
charges for the subsequent smelting process and decrease the 
amount of fluedust formed. 

That a circular furnace having the form of an inverted cone 
cannot fully possess the same advantages is evident, but whether 
the enlargement between the tuyeres and the throat of the furnace 
be sudden or gradual, the relative areas of the hearth and of the 
throat remain nearly the same — 1 : 2 or 2£. 

The shaft (c.c. Figures 90, 98-100, 102) is made of common 
brick c, and lined with fire-brick c' up to the feed floor. It 
rests on four supporting plates t, which used to be (Figures 84, 85, 
87, 88, 98) fastened to the capitals o\ of the j)illars k, but the 
effect was to loosen the pillars by the unequal expansion of brick- 
work and cast-iron plates, and thus endanger the safety of the shaft. 
To relieve the pressure upon the plates, brick arches ((/', Figure 96) 
were introduced, supporting the walls of the shaft and throwing 
the weight .upon the pillars. To counteract the lateral thrust the 
lower part of the shaft is enclosed in cast-iron plates (Figures 84, 
85, 96, 98) firmly bolted together, in addition to which the plates 
t (Figure 98) sometimes have flanges t'. In the latest furnaces 
a set of three I-beams (n', Figures 89, 90, 93, 99, 100) on each 
side of the furnace form the support of the cast-iron plates t 
(Figures 99, 100), being firmly bolted to each other and screwed 
tightly to the capitals o' of the pillars. The cast-iron plates t are 
in no way fastened to the beams, but rest freely upon them. By 
this arrangement supporting plates and shaft can expand indepen- 
dently of one another without endangering the stability of the shaft. 
The plates, being supported by the I-beams, no longer need to be 
reinforced by the arches, but can bear the weight of the shaft safely. 

The walls of the modern furnaces are made very thick at the 
bottom in comparison with older ones (32-^-39 inches against 17| 
inches, Figures, 89, 90, 99, 100, against 87, 88), decreasing toward 
the feed floor, which causes a considerable saving of fuel. 




1 

• 












BLAST FURNACE 

COLORADO IRON CO. 




B^ig. lOl 





BLAST FURNACE, 

MONTANA SMELTING CO. 




■"■c^T *\ 



184 METALLURGY OF LEAD. 

The entire shaft is well braced with tie-rods d', secured in 
corner-irons c' (Figures 84, 85, 89, 90, 95, 96). 

d. Feed Holes and the Collecting of Fumes. — There are two 
general arrangements for feeding and carrying off fumes. The 
first is to have a feed hole on either side of the furnace m f (Figure 
102). The stack or chimney c (Figures 102, 103) is of brick, forming 
the continuation of the shaft ; it is contracted at the upper end to 
about 3^ feet square (inside measurement), and then passes through 
the roof. The top is closed by a swinging damper resting in the 
groove J", and can be opened from the feed floor by means of a 
damper rod. 

In small furnaces the feed holes are placed in the middle of the 
sides, in large ones nearer the front and back of the furnace re- 
spectively, i.e., not opposite each other, but diagonally. This 
makes it much easier to distribute the charge evenly and to bar 
down the wall accretions, dividing the furnace practically into two. 
The doors (Figures 108, 109, 110) are 5 or 6 feet high, in order 
that a man may be able to stand in them and direct the bar. 
During the run they are closed to from 18 to 24 inches by letting 
down a sheet-iron curtain e" that slides in a cast-iron frame 
(Figure 108) and is balanced by counter-weights. The sill of the 
feed floor is placed about one foot above the floor level, so that the 
feeder cannot simply shovel the charges into the furnace, but, being 
obliged to raise the shovel each time, will be more likely to dis- 
tribute the charge evenly. 

The fumes are drawn off near the top of the stack by a circular 
sheet-iron flue v f which passes at a steep angle into the dust 
chamber. In the flue is the damper (Figures 102, 104, 105, 106) 
to regulate the draught, which is kept just strong enough to pre- 
vent the gases from passing out through the feed holes. In blow- 
ing in or blowing out, the damper in the flue is closed and the one 
on the top of the stack thrown open. 

The other arrangement is to feed the furnace from the top 
(Figures 87-91, 96-98). This is covered in part by cast-iron plates 
w' (Figures 91, 98), leaving an opening through which the 
charge is introduced. Formerly the Pfort method (Figures 87, 
88) of carrying off the gases was in general use. It consists in 
suspending from the cast-iron top-plates w' an iron curtain w 9 so 
as to leave room between it and the walls of the furnace for the 
gases, whence they pass off through a flue v into the dust cham- 
ber (Figure 81), the charge filling the inside of the curtain up to 



SMELTING IN THE BLAST-EURNACE. 185 

the feed floor. While this arrangement proved very satisfactory 
in a good many ways, it had one great disadvantage, that it 
lengthened the time required for barring down wall accretions, 
because it was necessary to remove the curtain before and to put 
it back again afterward. For this reason the curtain has been al- 
most universally discarded and the gases are simply drawn off by 
a flue v (Figures 96, 98) at the back of the furnace, or better 
by two on the sides. To avoid sucking in air or letting out fumes, 
the feed opening m' in the cast-iron top plate w' (Figures 90, 91, 
98) is made rather small ; it is, however, large enough for the 
feeder to be able to spread his charge in any way that may be 
necessary, and to reach any part of the side walls with a bar when 
cutting out wall accretions. The flue leading to the dust chamber, 
which used to be sheet iron (Figure 88), is now commonly built of 
brick (Figure 96). It rests on heavy rails and is thoroughly bound 
with buckstays and tie-rods. 

Formerly a "telescope stack" (z, Figure 81, and J', Figures 96 
and 98) was suspended by chains W over every furnace. It is a 
sheet-iron pipe reaching through the roof and balanced by counter- 
weights V. Its lower part is enlarged to the oblong form of the 
feed opening and has a small feed door m f on either side. The 
stack is lowered when the furnace smoke is not sufficiently drawn 
off by the flue, or when the furnace is blown in or out, to carry off 
the gases into the open air. 

v At present, large smelting plants have (say) two of these 
sheet-iron stacks suspended from a traveller, to be used in case of 
necessity, and in some instances this stack has been thrown off en- 
tirely (Figures 89-95). Instead, above each furnace is suspended 
a | -inch cast-iron plate sufficiently large to close the feed opening 
m\ It is lowered when the furnace is being blown out, to prevent 
the fumes from passing to the feed floor; the joint is made air-tight 
by spreading moistened fine ore over it. 

Another manner of collecting the waste gases with open- 
mouthed furnaces may be mentioned here. This is the Darby 
tube, ;i wrought-iron pipe of small area, in comparison with that of 
the throat of the furnace (280 square inches as against 3,207). It 
is hung in the middle from the girders, some distance (5 feel -J, 1 
inches) down into the furnace. This tube is used in the Upper 

Hartz Mountains, where fine galena concentrates are smelted raw. 
The advantages claimed for it are thai the charge is less Liable t<> 
pack toward the centre, and that the gases, being drawn off there, 



186 METALLURGY OF LEAD. 

are prevented from rushing up at the sides, and penetrate the 
charge more evenly. 

The idea of closing the throat of the furnace with cup and cone 
and feeding automatically has been often suggested, e.g., by Halm l 
but not tried until lately; the results of the experiments have y 
however, not been made public. 

e. Hearth with Arents > Automatic Tap. — The bottom of the 
hearth is formed by a bed plate of boiler iron y' (Figures 89, 90, 
96, 98-100), which is to prevent any lead from percolating down- 
ward. It is placed on the foundation, as indicated in § 59, b. Care 
must be taken to have its centre coincide with that of the shaft by 
dropping a plumb-line from the feed floor. The bed plate some- 
times has an angle-iron rim enclosing the bottom course of brick 
(Figures 89, 90, 99, 100); sometimes it reaches beyond the castings 
(Figures 96 and 98), enclosing the hearth, which rests upon it. 

These castings h' (Figures 86, 89, 90, 92, 96-101), reaching to 
the top of the hearth, have been (and still are) a great trouble, as 
they are very liable to crack. At first they were made 1 inch 
thick; later the front and back plates were strengthened by ribs i 
(Figures 89, 90, 92, 96-101); then the side plates, castings, and ribs 
have been made thicker and the bevelled corners have been fastened 
together by special tie-rods b' (Figures 89, 92). Still there is 
danger of their cracking, so that in some furnaces a wrought-iron 
band v' (Figure 90) is screwed to the sides to hold the casting to- 
gether. It would seem as if making the outer wall of the hearth 
oval and enclosing it with a f-inch wrought-iron plate, as is done 
in the modern large iron blast-furnaces, would be the way to solve 
the difficulty. 

To the casting on the front is fastened with bolts the slag spout 
a (Figures 89, 91, 96, 97, 99, 101), and to the side the lead spout a' 
(Figures 98, 100, 101), if the well is confined within the hearth 
plates, as is now usual. It is not often that the slag is tapped al- 
ternately from the front and the back of the furnace, requiring two 
slag spouts (Figures 89, 92, 96, 97), and that the lead is removed 
from the two sides (Figure 98). Two slag taps have been used at 
some furnaces to counteract the forming of a crust at the back of 
the furnace, where it usually begins growing toward the front, 
from which it cannot easily be reached. If the slag is tapped 
from both front and back, the danger of crusting at the back is in 

1 "Mineral Resources of the United States," 1882, p. 343. 



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SMELTING IN THE BLAST-FURNACE. 187 

part at least avoided, and if an obstruction forms there, it is easily 
removed. Having two lead-wells is simply a waste of heat, for 
when one becomes clogged up, the other will also. 

The hearth walls and bottom are of fire-brick. They usually 
rest on the bed plate (Figures 87, 88, 90, 98). Sometimes on the 
bed plate and below the crucible a 6-inch layer of ground brick and 
raw clay (3 : 2, by volume) is beaten down firmly in the form of an 
inverted arch (Figures 99, 100) on which the bricks forming the 
"bottom proper are placed. In building the side walls it is better 
not to place the bricks in direct contact with the castings (Figures 
■87, 88, 98-100,) but to leave a small space of (say) 2\ inches and 
tamp it out with brasque (equal volumes of ground coke and clay) 
while the bricks are being placed (Figure 90). In this way the 
crucible, when it expands, will simply pack the brasque tighter, 
and thus relieve the castings from at least part of the strain. 

Arents' Automatic Tap, or siphon tap, which forms part of the 
side wall, consists of an inclined channel, the siphon d (Figures 
87, 90, 98, 100), 3 or 4 inches square, running from the lowest part 
of the crucible wall inside to the top on the outside, where it is en- 
larged into a dish-shaped basin — the lead-well c (Figures 87, 90, 
97, 98, 100, 101), the length, of course, depending on the depth of 
the crucible, which varies from 22 to 30 inches. The tap is usually 
in the middle of one of the sides (Figures 96, 97) although some- 
times placed nearer the front (Figure 101); while the furnace is 
running, the crucible remains nearly full of lead, that in the auto- 
matic tap standing a little higher on account of the pressure of the 
blast. From the well the lead, as fast as it is made in the furnace, is 
either ladled into moulds or it overflows into the cooling-pot, or it 
is periodically tapped into it, and is thence ladled into moulds. 
When first used, 1 Arents' tap consisted of a 3-inch wrought-iron 
pipe which terminated a foot or more below the upper rim of a 
sheet-iron cylindrical shell rammed full of fire-clay and bolted to 
the casting, the well being afterward cut out. The wrought-iron 
pipe has been universally abandoned. The remains of the sheet- 
iron cylinder are still found in some instances in a half-cylinder 
that is bolted to the casting (Figures 86, 87). In most furnaces 
to-day the lead-well is enclosed in the crucible wall (Figures 90, 
96, 98, loo, ioi). The advantages of this improvement are that 
the siphon is shortened and the lead is kept hotter. In smelting 

1 Hahn, Eilers, Raymond, Trans. A. I. M. E., i., p. 108. 



188 METALLURGY OF LEAD. 

charges that run high in lead, the lead in the crucible is frequently 
exchanged, and a slight loss in heat does not make itself felt; the 
siphon can be long, and the well, not being close to the tuyere- 
pipes, will leave these cool. With charges low in lead, all h 
heat must be avoided. The increased thickness of the side wall, 
necessitated by the enclosed lead-well, is altogether an advantage, 
because the loss of heat is diminished. In any case, the side walls 
ought never to be less than 2£ inches thick. 

The advantages of the automatic tap are many. Without it 
the lead and matte are tapped from the bottom of the crucible. 
To do this the blast is shut off, the blast-pipes are removed, and 
the tap-hole is opened, which is often done with difficulty. Lead 
and matte run out into a shallow tapping-kettle, and the moment 
the slag appears the opening must be closed with a stopper of 
brasque or clay. Then the crucible is cleared by inserting through 
the fore-hearth (if it has one) a curved iron bar, the tuyeres are 
cleaned with iron rods, the blast-pipes are put in place, the blast is 
then turned on again slowly, and smelting is resumed. This stop- 
page takes considerable time, and therefore cools the furnace. 
Then into the crucible (now free from lead and matte, although it 
retains some fuel) falls an equivalent amount of half-melted charge, 
which has gradually to be lifted up by fresh lead and matte, when 
tapping begins again. These half-melted masses thus have a 
chance to adhere to the bottom of the crucible, and are very apt to 
be the beginning of a bottom crust. If this has once started, it is 
nearly sure to grow and gradually freeze up the furnace. 

With Arents' tap there is no stoppage when the lead is re- 
moved from the lead-well or when matte and speise are tapped 
with the slag at intervals into a slag-pot, where they settle out 
according to their specific gravities. The furnace therefore runs 
more regularly, and the first formation of slag or matte accretions at 
the bottom is prevented, as the crucible is always filled with lead. 

The claim originally made that the lead from the automatic tap 
is purer than that from the tapping-kettle will hardly be main- 
tained to-day. It was based on the theory that the metal, being 
taken continuously in small quantities from the bottom of the 
crucible, would be purer, as the heaviest, i.e., the purest lead, would 
gather there, and the impurities would float to the surface, to be 
taken up by matte and slag. This presupposes that the lead in the 
crucible is sufficiently undisturbed to permit liquation. The facts 
are that there is a constant current in the lead which prevents this 



SMELTING IN THE BLAST-FURNACE. 189 

separation of dross and lead. The dross is disseminated through the 
lead; some reaches the surface and is taken up by matte and slag; 
but a large part of it rises in the siphon and collects in the well. 
Thus the bars from the lead-well are often less pure than those 
from the tapping-hearth, as with the latter the dross adheres to 
the cake of matte floating on the still liquid lead. However, by 
skimming the dross from the well or cooling-pot, clean bars can be 
obtained and the dross returned at once to the charge, thus in- 
volving little loss in metal. With the tapping-hearth, the dross 
adhering to the matte must undergo all the operations with it, and 
thus much lead and silver are lost. The automatic tap, therefore, 
insures a considerable saving in metal. 

There is one case where tapping from the bottom is to be pre- 
ferred to the automatic tap. It is in smelting coppery ores — as an 
alloy of lead and copper separates Out from the lead, adheres to the 
bottom of the crucible, and grows upward, filling it (see § 54, i). 
The trouble is remedied by adding sufficient sulphur in some form 
or other to form matte. As soon as this runs 12 per cent, copper 
the difficulty begins again to make itself felt. In concentrating 
lead matte from 12 per cent, copper upward, the automatic tap is 
out of place, and the ordinary, copper furnace with internal cruci- 
ble is used. 

f. Water-Jackets.— These (B, Figures 89, 90, 92, 96, 98) are water- 
cooled iron shells that enclose the smelting zone of the furnace to 
protect it from the corrosion of the slag. Since about 1873 they 
have come into more general use, and have now entirely replaced 
the brick walls at the region of the tuyeres, whenever there is suf- 
ficient water to warrant their use. Only where this is very difficult 
to obtain, sandstone, fire-brick, or other refractory material is re- 
verted to. 

Quite a discussion arose in 1885-86 l as to the date of the in- 
vention of the water-jacket furnace and the inventor. The writer 
has failed to find any reference to water-jackets in the treatises of 
Karsten 2 and Scheerer. 3 The earliest mention of their use is made 

1 Engineering and Mining Journal, 1885 : July 25 (Harnickel, R, Rol- 
Ker) ; August 1 (Courtis); August 15 and 29 (Kleinschmidt) ; August 22 
(Editor) ; September 12 and 26 (Williams) ; October 10 (Hahn) ; October 24 
(Arents) ; October 31 (Douglas) ; November 7 (Courtis, Daggett) ; November 
14 (Curtis) ; November 29 (Kleinschmidt) ; 188G, January 2 (Tew). 

2 "System der Metallurgie," Berlin, 1832 ; and " Handbuch der Eisen- 
huttenkunde," Berlin, 1841. 

3 "Lehrbuch der Metallurgie," Brunswick, 1840-53. 



190 METALLURGY OF LEAD. 

by Overman, 1 who describes and illustrates a refinery furnace, the 
sides of which consisted of water-cooled cast-iron shells, through 
which water-cooled tuyere nozzles protruded into the furnace. 
Douglas 2 says that J. Williams built near Drontheim, Norway, in 
1852 "sectional water-jacket furnaces consisting of a circle of long, 
narrow water-backs, perforated by tuyere-holes." About the year 
1S65 the same J. Williams erected a number of water-jacket 
blast-furnaces at Houghton, Lake Superior. According to Arents, 3 
N. Haskell built in 1865 a water-jacket furnace in California. 
Kerl, 4 in describing the improvements made in smelting in the 
Hartz Mountains, records the introduction in 1864 of " water - 
blocks to cool the hearth and to serve as a support for the water- 
cooled tuyere nozzles," but these had been used in refinery furnaces 
for a very long time, 5 and are not to be confounded with water- 
jackets. The latter never were and are not to-day in use in the Hartz 
Mountains. Courtis, who made a drawing of the Pilz furnace at 
Freiberg in 1886, says that the tuyeres and the front of the furnace 
were water-cooled. Water-jackets have been introduced at Frei- 
berg and Pribram only since they became common in this country; 
they were used in the Saint-Louis Smelting Works near Marseilles 
before 1878. 6 Spray-jackets were used at La Pise as early as 
1862. 7 

The water-cooled tuyere nozzles, which resist the action of heat 
and slag so well, while the brick walls are eaten out, appear to 
have suggested to different persons the idea of extending the 
water-cooled iron surface, and thus caused the construction of the 
water-jacket. This would give several men the credit of having 
invented it. 

The water-jackets (F, Figures 90, 98 ; F y Figures 90, 100) are 
placed on top of the hearth-walls, forming their continuation on the 
inside. Their height has varied from 2 to 4 feet ; 3 feet 6 inches 
is an ordinary measure. They thus reach from the top of the 
hearth to within about 12 inches of the cast-iron carrier-plate (t) 
or the I-beams (>*') which support the shaft. The centre of the 

1 " Treatise on Metallurgy," New York, 1852, p. 556. 

2 " Mineral Resources of the United States," 1882, p. 268. 

3 Berg- und Hiittenmcinnische Zeitung, 1866, p. 316. 

4 Ibid., 1867, pp. 6 and 47. 

5 Percy, ''Metallurgy of Iron and Steel," London, 1864, pp. 584 and 625. 

6 Gruner, " Traite de metallurgies' Paris, 1873-78, vol. ii., p. 391. 

7 Gruner, Annates des mines, 1868, xiii., p. 364. 



SMELTING IN THE BLAST-FURNACE. 191 

tuyeres is placed 10 inches higher than the bottom of the jackets, 
and from 8 to 10 inches above this, begins the bosh, the amount of 
which varies from 6 to 10 inches. 

Jackets are made of cast iron and wrought iron. 

Cast-iron jackets (J5J Figures 84, 85, 87-89, 90, 92, 96-98) are 
generally 6 inches thick, the sides being of |- or f -inch iron. Each 
jacket has a special water-feeder/, which begins 8 or 10 inches 
above the centre of the tuyere and runs from 3 to 4 inches above 
the top of the jackets, extending outward about 4 inches. This 
insures the complete filling of the jacket with water. As the top 
of the feeder is closed only by a lid, tools can be introduced to 
scrape off scale. The feeder was formerly cast in one piece with 
the jacket, but now is a separate casting, which is fastened on 
with screws or bolts. At first there was no opening at the lower 
end of the jacket (Figures 84, 85, 96) to remove mud or scale that 
had collected ; now there is usually a hand-hole {q', Figures 
89, 90) for this purpose, and thus the life of the jacket is much 
prolonged. The tuyeres (o, Figures 84, 85) always used to be 
at the junction of two jackets, each having a semicircular re- 
cess. This was for fear that an opening through the centre of 
the jacket might weaken it ; but it has not proved to be the 
case, and the tuyeres are now (Figures 89, 92, 96, 97) generally 
made in that way. Tuyeres between jackets are sure to cause 
considerable leakage of air. When a furnace is new, and the two 
semicircular recesses are just opposite each other, and a brass noz- 
zle (]/, Figure 98) is inserted, which receives the galvanized iron 
blast-pipe, this does not at once occur but is sure to come later. 
Not less than six different kinds of jackets (Figures 84-87) were 
formerly used in each large furnace ; now the number has been 
reduced to three (Figures 92, 97), two (A and B) on the front one, 
and one (E) on the side, the opposite ones being duplicates. To 
these must be added the tapping-jacket (n, Figures 84, 90, 99). 
This reduction in number is made possible by giving the front of 
the furnace the same construction as the back, the opening re- 
quired at the front being simply bricked up at the back. The 
length of the front jackets varies somewhat in different furnaces. 
Sometimes they reach to within 10 inches of the top of the cruci- 
ble (Figures 99, 111), thus leaving open a 10-inch breast, which 
runs across the entire width of the furnace. It is usually closed 
by two small brick pillars (9 by 4^ inches) and three balls of clay, 
one in the middle, and one on the outer side of each pillar. In the 



192 METALLURGY OF LEAD. 

central clay -ball is placed the slag-tap. There is nothing in this 
arrangement to prevent a tapping-jacket from being put in, 
should the clay be eaten out by the slag. A second plan is shown 
in Figure 84. Here the front jackets reach at the sides down to 
the top of the crucible, and leave in the centre an opening, the up- 
per part of which is closed by a separate jacket n, and the lower, 
the breast /, by a ball of clay, in which is placed the slag-tap. 
This plan is not a common one. A third modification (Figures 90, 
98) is to fill the open place in front with a tapping-jacket to within 
2\ inches of the top. This space is left for convenience in taking 
out the tapping-jacket, and is closed by brick. The jacket (/<, 
Figure 99), 26 by 14 inches, and 3^ inches deep, has, 6-J inches 
above its lower edge, a tap-hole m, which is 2-J inches in diameter, 
and widens, after entering the jacket to the depth of one inch, to 5 
inches on the inner side. The lower edge of the jacket is placed 4 
inches beneath the upper edge of the crucible castings, and thus 
prevents at the front that leakage of lead from the crucible which 
is otherwise so difficult to stop. The last arrangement is the one 
now in general use at large smelting works. Having large ore- 
beds from which to make up the charges there is little probability 
of trouble in the furnace, and therefore a small opening closed by 
the tapping-jacket is sufficient for all practical purposes. At small 
smelters, where the charges are changed often, it is probably better 
to have the front jackets 10 inches smaller than the side jackets, 
in order to give room for working in the crucible, should it prove 
necessary. 

The water-jackets have been joined in various ways by using 
wedges, screws, bands, etc. Now they are simply bolted to each 
other near the top and bottom, the bolts passing through lugs 
(y, Figures 89, 92, 96, 97) cast in the jackets. 

Wrought-iron jackets have in many instances replaced those of 
cast iron. AVith circular furnaces they are exclusively used. 
They have no bosh, and are usually in two parts. They seldom 
have special w T ater-feeders, like the cast-iron jackets, the water 
inlet pipe being usually near the bottom and the outlet pipe near 
the top. AVith this arrangement, it is important, if a complete 
filling of the jackets with water is to be made possible, to have 
one or two small pieces of pipe protrude upward and outward from 
the top of the jackets, and to have the water outlet pipe also bent 
upwards that it may discharge above the top of the jackets. 

Oblong furnaces have only four wrought-iron jackets, one 



SMELTING IN THE BLAST- FURNACE. 193 

on each side. Two kinds of jackets are in use. One is shown in 
Figures 99-101, and 112-115. The jacket is of the usual height ; 
the inner wall forms a straight line slanting outward ; the outer 
wall has a greater slant ; thus the water-space grows wider toward 
the top, where the water outlets i are riveted to the outer wall, the 
inlets being at k" . The walls are stiffened by stay-bolts. The 
mud accumulating in the jackets is removed through the hand- 
holes q'. The details of the tuyere-box a" are discussed further 
on. The other form of jacket is like the cast iron one. The 
walls are vertical and parallel from the base to the bosh, and then 
slant outward ; cast-iron water-feeders are bolted to the outer 
wall. If cast-iron and wrought-iron jackets be compared, there is 
no question that the latter last longer than the former, and that, al- 
though much more expensive at first, they are cheaper in the end. 
There may, however, be exceptions to this rule. If the water 
that is to cool the jackets be muddy or hard, and thus liable to form 
scale, a wrought-iron jacket will not last so much longer than one 
of cast iron. Again, if a foundry be not too distant from the 
smelter, the value of the leaking cast-iron jacket as old iron will 
contribute considerably toward the cost of a new one. 

Water-jackets made of cast steel are strongly recommended 
by Terhune, 1 who introduced them at the Hanauer Works, where 
they gave great satisfaction. The walls are § inch thick. 

The cooling-water for the jackets is drawn from a wooden tank, 
the bottom of which should be at some distance above the water 
inlet, in order that there may be some pressure, as an extra amount 
of water is often needed in blowing in or blowing out. The main 
delivery-pipe starts from the water-tank and runs along the front 
or back of a row of furnaces. From it branch off separate supply- 
pipes (.'•/ Figure 85), each of which ends in a pipe g (Figures 84, 
85, 89, 00, 92, 96, 98-100), surrounding its own furnace. This 
supplies the small feed-pipes h, which deliver the water into the 
top of the feeders, the flow being regulated by a valve. The cold 
water, entering 1 lie jacket at the top, sinks down slowly and pushes 
Upward the hot water, which runs off through the small pipe 
I below the inlet. This, the common arrangement, accomplishes 
on the whole its purpose, but if the temperatures of a jacket he 
compared at the top and the bottom, il will be found that the 
bottom is always hotter. There are two ways of equalizing this. 

1 Trams, a. j. m. a.. Kvii.,p. 131. 



194 METALLURGY OF LEAD. 

One is to attach a rubber hose to the feed-pipe A, thus letting the 
cool water come in contact with the hot water at about the middle 
of the jacket. The other is to have an extra supply-pipe through 
which a small stream of water runs in near the bottom of the 
jacket. If these two methods do not succeed in cooling the lower 
part of the jacket, it shows that it contains mud or scale, and re- 
quires cleaning, if it is not soon to burn through. The hot water 
from the jackets is discharged into a galvanized-iron trough (Figures 
84, 85, 96, 98) j, surrounding the furnace, from which it passes off 
through a cast-iron stand-pipe It (Figures 85, 96) into a main under- 
ground. The troughs are often in the way when the furnace is 
running, and are very inconvenient sometimes, as, for example, 
when a cracked jacket has to be exchanged. To remedy this, the 
water is sometimes carried away from the jackets by long pieces 
of gas-pipe, terminating in the funnels of stand-pipes placed close 
to the supporting pillars and connected with the main underground. 
There are two or four of these stand-pipes. Thus, if a jacket has 
to be exchanged, it is only necessary to turn off the water-supply 
from it instead of from the whole side, as is ordinarily done, the 
trough being also removed. The amount of water required to cool 
the jackets varies with the size of the furnace and the slag that is 
being made. A furnace 36 by 92 inches at the tuyeres, making 
a siliceous calcareous slag, requires under normal conditions 11 
gallons of water per minute. This is a good average figure. For 
blowing in or blowing out one must be prepared to use double this 
amount. It is sometimes said that cast-iron jackets require less 
water than wrought-iron ones. This remains to be proved. 

g. Blast. — The machines that furnish the blast belong to the 
class of rotary positive pressure-blowers. Fans are used very rare- 
ly. A comparison of the two has been made by Howe. 1 The 
pressure-blowers at present in general use are the Baker and Root. 
The former was at one time almost exclusively used, but the latter is 
gaining ground, and is now found in a good many works. Its ad- 
vantage is that it allows a greater difference in speed. Take, for 
example, a No. 7 Baker blower, giving a displacement of 60 cubic 
feet per revolution. Its greatest allowable variation in speed per 
minute is 35 revolutions ; the No. 7 Root with 65 cubic feet dis- 
placement allows 100 revolutions. If a higher pressure is wanted 
in a furnace, the Root can therefore be more easily speeded up 

1 Trans. A. I. M. E., x., p. 482. 



SMELTING IN THE BLAST-FURNACE. 195 

than the Baker. A furnace of from 36 X 60 to 30 X 84 inches 
will require a Baker or Root No. 6 ; a furnace varying from 36 X 
100 to 36 X 120 inches, a Baker or Root No. 7. These sizes give 
more blast than is actually necessary, but it is advisable to have a 
slight excess. 

There are two ways of supplying the blast to a number of 
furnaces. Each furnace can have its own blower or several 
blowers deliver the compressed air into a blast-main, from which 
the single furnaces are supplied. The second method is the one 
generally accepted, as the plant is cheaper, the dail}^ attendance 
easier, and the repair smaller. The pressure in the single furnaces 
is regulated by a gate in the branch-pipe. All blast-pipes are made 
of galvanized iron. The main blast-pipe or induction-pipe (s, 
Figure 81), which receives the wind from several blowers, has a 
diameter that is from-g- to § larger than the combined outlets of the 
blowers. It usually runs along back of the furnaces near the dust 
chamber, and is suspended eight or more feet above the furnace 
floor in a wooden frame. It has safety-valves, and is closed at 
both ends by blast-gates. From it branch-pipes, each with its own 
gate, furnish the blast to the bustle-pipes of the single furnaces 
(Figures 84, 85, 90, 96, 98-100). Back of the gate each branch- 
pipe has an opening, with a thimble, to be connected with the 
pressure-gauge. 

The pressure of the blast is measured by different quicksilver 
or water gauges, 2 inches quicksilver or 28 inches of water equalling 
1 pound or 16 ounces pressure per square inch. At some works 
colored glycerine is used instead of quicksilver or water. 

From the bustle-pipe thimbles pass downward to be connected 
by the wind-bags q with the tuyere-pipes p. The wind-bags are 
of closely-woven canvas that has been soaked in water-glass or 
alum to prevent it from readily catching fire from a spark. The 
number, size, and form of tuyeres and tuyere-pipes vary a great 
deal. Iluhn 1 reckons one tuyere with a 3-inch opening as being 
sufficient for 2 square feel hearth area. Thus a 3 X 5 foot furnace 
would require seven tuyeres; one that is 3 xC| feel, nine tuyeres. 
They are so distributed that one tuyere enters at the back and 
three or four are inserted symmetrically on each side. At sonic 
works the diameter of the tuyeres lias been reducced to 2^ incites ; 
at others it lias been increased to 3j inches, but these figures are 
not common. 

1 " Mineral Resources of the United States," 1882, p. 336. 



19G METALLURGY OF LEAD. 

The tuyere at the back has been lately abolished at many 
works, while the tuyere at the front (Figures 84, 85) was discarded 
many years since as being very much in the way and of no special 
advantage. Blowing only from the sides has the effect of making 
the breast hard; blowing from the sides and ends of chilling in the 
centre. The former difficulty can be avoided by having the last 
tuyeres close to the ends, and also by the manner of feeding; the 
latter is not easily remedied. Hence it is advisable to leave out 
the end tuyeres. 

The ordinary blast-pipe is made of galvanized iron. The hori- 
zontal arm, varying from 2 to 14 inches in length (Figures 84 and 
90), is either slightly conical and fits into a brass nozzle p' (Figure 
98) inserted into the tuyere-hole, or it is cylindrical and is soldered 
to the nozzle. The elbow which joins the other end to the wind- 
bag has a brass nipple soldered to it, which forms the eye or peep- 
hole s' (Figures 90, 96, 98). This is closed either by a slide or a 
cap having a glass or mica plate in the centre, or simply by a 
wooden plug. In the centre of the plug is left a small opening, 
the size of a pencil, to be closed by a small piece of wood. This 
has to be removed to observe the condition of the tuyere. To keep 
the blast-pipe in its normal position and to thus prevent it from de- 
livering the blast upward, which is its natural tendency, an iron 
band, hooked by means of two springs to the jacket, is passed 
around the elbow, or an iron loop is soldered to its inner side, by 
means of which it is hooked to the jacket. In order to close the 
tuyere-hole, when the pipe has been temporarily removed, a tuyere- 
cup has in many works replaced the ball of clay commonly used. 
It is made of galvanized iron, has the form of a nozzle, and is 
closed at the back, to which is soldered a handle. 

In this connection may be mentioned Werner's adjustable 
tuyere-pipe. 1 It consists of a cast-iron pipe to which a cast-iron 
elbow is fastened on the upper side by means of a hinge. The pipe 
is hooked with three springs to the jacket, and, working in a ball- 
and-socket joint, can be turned in any direction, and thus deliver 
the blast wherever desired, the springs keeping the pipe in position 
and very little air being lost by leakage. Other advantages are 
that should any slag run into the pipe it will not damage it and 
can be easily removed by raising the elbow. If the furnace is to 
be shut down for a short time, which usually necessitates the re- 

1 Emmons, "Geology and Mining Industry of Leadville," p. 682. 



198 



METALLURGY OF LEAD. 



nioval of the pipes, a piece of thin card-board inserted between the 
flanges at junction of pipe and elbow is sufficient. The writer has 
used this tuyere-pipe, and while there is not as much advantage as 
is claimed in being able to set the blast in any direction, its other 
features make it a good apparatus, preferable to the ordinary pipe 
of galvanized iron. 

In the ordinary blast-pipe, even with the best care, a leakage 
of air cannot be prevented. This has led some works to adopt a 
cast-iron elbow or tuyere-box, which is fastened by an air-tight 
joint to the jackets. The constructions of Eilers, Murray, and 
Devereux may serve as examples. 

Figures 117--120 represent the automatic tuyere-valve designed 




ENDVIEWEF. 

Fig. 134 



THE MURRAY TUYERE-BOX. 

by Eilers. It consists of a cast-iron box a", which is fastened with 
cap-screws to the jackets, as shown in Figures 99, 100, 112-115. 
On the upper side is a cast-iron nipple jp, to which the wind-bag is 
fastened ; on the lower side a 3-inch opening through which chilled 
slag can be removed, to be closed with a plug. At the rear end are 
two openings : I", the peep- and poking-hole, and m\ the outlet for 
the back pressure of the furnace gases when the blast is shut off. 
When the furnace is running, the valve n", swinging on the cross- 
pin o" and balanced outside of the box by the movable weight p' 1 \ 
is pressed by the blast entering through p against in" and closes 
this; as soon as the blast is taken off, the weight/)" turns the 
valve, closes p, and opens m", thus preventing the gases in the 




LONGITUDINAL SECTIC 





.T10HOJ 



SMELTING IN THE BLAST-FURNACE. 



199- 



furnace from entering the blast-pipe and furnishing them an outlet 
into the open air. 

Murray's tuyere-box is shown in Figures 121-124. Figure 121 
shows an obtuse-angle elbow. The opening a through which the 
blast enters the jacket is rather large — 5 inches in diameter. Op- 
posite is the poking-hole b closed by the cap c, which contains the 




Fig. 127 

F 



-0% 



COMMON AXLE. 
2-w.l. 



COMMON SLAG-POT. 




eye-hole. Between the tuyere-pipe d and the cast-iron nipple e, 
over which passes the wind-bag, is the wind-gate/. Of special 
interest is the drop bottom (j of the belly-pipe //, shown in Figures 
122 and 123. It is held in place by the hinged bolts i and the 
crank-nuts j. The drop bottom has a fusible plug /■, a thin disc of 
lead, which will melt as soon as any slag enters the belly-pipe, and 
thus call immediate attention to the accident. The elbow is 



200 



METALLURGY OF LEAD. 



fastened to the jacket by slipping it over four threaded bolts m 
that have been screwed into the jacket, and then tightening it with 



nuts. 



The characteristic of the Devereux tuyere-box l is that the blast 




Fig. 129 



k&- 



-SC- 



ROLLER BEARING AXLE, 
2-M.S. 




:FM S .130 



a 



— 3 Ke— 

ROLLER. 

18-M.S. 



Ki g . 131 




SLAG-POT WITH ROLLER BEARINGS. 



can be made to play in different directions. The tuyere-hole is 
lined with a cylindrical bronze tube in which can be revolved an 
iron sleeve having a diagonal bore. By turning this sleeve the 

1 Patent No. 318,604, May 26, 1885. 



SMELTING IN THE BLAST-FURNACE. 



201 



blast can be directed up and down to the right and left; it cannot, 
however, be directed centrally. 

h. Slay-Pots. — Twelve ordinary slag-pots, 24 inches in diam- 
eter and 15 inches deep, are sufficient for a 33- by 100-inch fur- 
nace. Considerable improvements have been made in the construc- 
tion, the aim being to make them light without diminishing their 
strength. 

Figures 125 to 128 l show the usual paraboloid form of a slag- 
pot with the ordinary compression-spoke wheel, the spokes being 
wrought iron, the axle machine steel. The length of the handle 
is 5 feet, the height of the cross-piece above the ground 2 feet 8 
inches. 

In order to lessen the friction at the hub, roller bearings (a, 
Figures 128-131 *) have been introduced with satisfactory results. 




SECTION/ L SLAG-POT. 

Terhune, 2 Figures 132 and 133, has made the bottoms of slag- 
lots movable, so that they can be replaced when corroded or 
cracked. 

When matte and slag are being tapped, there is danger that 
they may not separate well. To assist their doing so several slag- 
pots have been constructed. 

The Iles-Keiper pot 3 is a large overflow-pot which retains the 
matte, while the slag runs over through a spout in the side into an 
ordinary slag-pot. To prevent the melted mass from solidifying 
the pot has a cast-iron cover. Overflow-pots have been used for a 

1 Taken from a drawing of Messrs. Fraser and Chalmers, Chicago, III. 

■ Trans. A. I. M. E., xv., p. 92. 

8 Patent No. 335,224, February 2, 1886. 



202 METALLURGY OF LEAD. 

good many years. At some works the slag is allowed to harden 
in them from the surface down to a depth of several inches, the in- 
let and overflow for the slag being kept open. Thus the hardened 
slag, taking the place of the iron cover, prevents the liquid slag 
below from cooling, and promotes a good separation of matte and 



Werner 1 has patented a slag-pot which permits the separate 
pouring off of the bulk of the slag. On the rim of the bowl op- 
posite the handle a segmental cover is pivoted, which is provided 
with a spout for the discharge of the slag. When the pot is filled, 
the slag is allowed partly to solidify, is then broken, the pot tilted, 
and a certain amount of the still liquid slag is allowed to run out, 
the remainder solidifying with the matte and speise, and being 
removed with them. The writer has never seen this pot in use. 

A third method, and the one in common use for preventing a 
loss of metal by shots of matte adhering to slag, is to allow the 
matte to settle in a catch-pot and then to tap the still liquid slag 
above the level of the matte. Thus the matte in the bottom of 
the pots remains undisturbed, and the shell of chilled slag (that en- 
closed the liquid slag) is recovered and smelted over again, as it is the 
only part of the entire slag that is liable to be rich. Several patents * 
have been taken out for different applications of this method. 
The catch-pot of Murray (not patented), shown in Figures 134-139, 
will serve to illustrate the third method. The pot has a cast-iron 
bowl of the usual paraboloid form, but 3^ inches above the bottom 
is the tap-hole a, through which the liquid slag is discharged after 
the matte has settled out. To prevent injury to the pot at the tap- 
hole, when this is opened with a steel bar, the casting is made thicker 
by a ring b, ending in a rib c which reaches the top. 

In this pot is seen a third class of wheel. The hub d has alter- 
nate spoke-sockets e and e', similar to those of the tension-spoke 
wheel. This wheel is more common than the ordinary compression- 
spoke wheel, as it is stronger. 

The Kesmith Dumping Car, Figures 140-144. 3 — The object of 
this car, with its two large tilting-pots A A, is to convey the waste 
slag from a number of catch-pots near the furnace building to the 
edge of the dump and to discharge it there. By this means the 

1 Patent No. 216,648, December 21, 1886. 

2 For instance, Devereux : No. 312,439, February 17, 1885 ; No. 335,114, 
February 2, 1886. 

3 Patent No. 388,708, August 28, 1888. 



SMELTING IN THE BLAST-FURNACE. 



203 



distance that the pots receiving the slag from the furnace have to 
be wheeled is shortened, and the disposal of the waste slag cheap- 




ened. Each tilting-pot has a capacity of V.:}s cubic feet, and holds 
about 1,280 pounds of slag. The car consists of a truck-frame a, with 



204 



METALLURGY OF LEAD. 



platform b, brake c, and railing d, by which the driver holds on. 
The frame carries the boxes e for the axles fof the wheels y. 
Two bridge-beams h, lying transversely across the frame «, serve as 
support for the frame h', which carries the central pin-socket ???, in 
which the pin n is made fast by the nut n'. The swinging frame 
consists of the channel-irons i i (held apart by the central blocks 
jj, in which are the swivel-eye and the end-blocks j' f) and the 
beams Jck (resting on the channel-irons); the latter have on their 



Fig. 144 




DOUBLE, SIDE DUMPING, 
SLAG POT. 



underside the bearings of the trunnions Hot the tilting-pots A A. 
These are pivoted out of centre, so that when in their normal posi- 
tion they may lean against the stop-pieces o. They are held in 
position by the pawls r, with disengaging handles r', which fasten 
into the teeth q of the projecting head or collar^ of the trunnions. 
This collar also has the holes s for the operating-bar t (Figure 
144), with which the pots are tilted. The weight complete of truck 
and pots is 5,000 ])Ounds. When in their normal position, the pots- 









OJ 

- 

- 








Trip r 







3T 









_f 



SMELTING IN THE BLAST-FURNACE. 205 

are placed as shown in Figures 140 and 141. When they are to be 
emptied, the frame iikk is swung 90 degrees on the swivel (Figure 
144), the pots are emptied together on both sides of the track; or, 
if the slag is to be discharged only on one side, the frame is re- 
turned to its normal place after one pot has been tilted and the 
other pot swung into position. 

§ 60. Chemistry of the Blast-Furnace — The reactions 

of a lead blast-furnace have never been fully studied. Guyard x 
has a very interesting theoretical discussion of the subject in his 
paper, " Argentiferous Lead-Smelting at Leadville," but measure- 
ments of temperatures and analyses of gases are wanting. Scher- 
tel 2 has measured the melting temperatures of the slags and an- 
alyzed the gases of the Freiberg blast-furnaces, where slagged 
ores are smelted. With the aid of these two writers, whose work 
will be used to supplement the more general outline, the sequence of 
processes in the lead blast-furnace can be approximately given. 

The main reactions that take place in the blast-furnace are re- 
duction and precipitation ; the latter has already been discussed 
(§§ 7 and 54, 1). The principal reducing agents 3 are carbon and 
carbonic oxide. 

a. Carbon acts on metallic oxides if its own affinity for oxy- 
gen is greater than that of the metal combined with the oxygen, 
and this affinity stands in direct proportion to the height of the 
temperature. The product of carbon and oxygen will be carbonic 
acid, if the carbon acts on readily reducible oxides (lead, copper). 
With oxides not easily reduced (iron, manganese) carbonic oxide 
is the principal product. 

2PbO+C = Pb 2 +C0 2 , 
Fe.O^+C = Fe x O„+CO, 
Fe^+yC = Fe x +yCO. 

b. Carbonic oxide is formed in the blast-furnace : first, by the 
direct combustion of carbon, 

C+O = CO, 
yC+Fe.O^ yCO+.rFe ; 

1 Emmons, "Geology and Mining- Industry of Leadville," p. 731. 

2 Wagner's . /ah rcsbn-ichte, 1880, p. 188. 

8 Balling : "Compendium der metallurgischen Chemie," p. 68; Wright, 
in Colliery Guardian, January 8, 1890, or Roberts-Austen, "An Introduc- 
tion to Metallurgy," Philadelphia, 1801, p. 194. 



206 METALLURGY OF LEAD. 

secondly, by carbonic acid being split into carbonic oxide and oxy- 
gen, 

co 2 = co+o, 

0O 8 +Fe x O lf . 1 = CO+Fe x O„ 
2/C0 2 +*Fe = 2/CO+Fe x O y ; 

thirdly, by carbonic acid combining with carbon, 

C0 8 +C = 2CO. 

The reducing power of carbonic oxide is on the whole favored 
by a high temperature. Its product of combustion is carbonic acid. 

PbO+CO = Pb+C0 2 , 
Fe.O.-f CO = F^O^+CO,, 
Fe^+yCO = icFe+yC0 2 . 

As carbonic acid begins to be decomposed into carbonic oxide and 
oxygen at 1,200° C, and cannot exist at 2,000° C, it follows that the 
reducing power of carbonic oxide is diminished as the temperature 
rises above 1,200° C. Hence the same mixture of carbonic oxide 
and carbonic acid may be a reducing agent at a low and an oxi- 
dizing agent at a high temperature. 

c. Other sources of carbonic acid in furnace gases are : 
First, the splitting of carbonic oxide into carbon and carbonic 
acid at a temperature of from 400 to 450° C. when in contact with 
iron oxide. 

2CO = C0 2 +C. 

Second, the decomposition of carbonates : lead carbonate, 170 
-200° C. ; magnesium carbonate, 650° C. ; ferrous carbonate, about 
800° C. ; calcium carbonate, 850° C. 

Third, its entrance into the blast-furnace in small quantities 
with the blast : 0.04 per cent, by volume. 

The action of sulphur as a reducing agent has already been 
discussed (§ 9). 

As has been seen, the carbon compounds in furnace gases are 
mainly carbonic acid and carbonic oxide. The oxygen of the 
air, on entering the furnace through the tuyeres, meets incandes- 
cent fuel, which is converted into carbonic acid or carbonic oxide 
according to the prevailing temperature. This depends on the 
melting-point of the slag. 

Freiberg lead-slags melt, according to Schertel, at 1,030° C. ; 



SMELTING IN THE BLAST-FURNACE. 



207 



other temperatures 1 given for lead slags are 1,220° C. and 1,273° 
C. ; Guyard assumes 1,200° C. 

In the products of combustion carbonic acid will therefore 
strongly prevail over carbonic oxide. In ascending, some of the 
carbonic acid will be reduced by carbon to carbonic oxide ; to this 
will be added that which is produced by carbon acting on oxides 
of iron. These reactions decrease very quickly as the ascending 
gases expand at the boshes and come in contact with the cooler 
parts of the charge ; here the percentage of carbonic acid again 
increases quickly through the reducing action of carbon and car- 
bonic oxide on readily reducible oxides. Thus the percentage of 
carbonic acid in furnace gases increases, with one slight interrup 
tion, with their distance from the tuyeres, and that of carbonic 
oxide correspondingly decreases. The amount of carbonic acid 
will be further increased by the decomposition of carbonates in the 
charge. 

Of the nineteen analyses of gases from Freiberg blast-furnaces 
made by Schertel, 2 a few, selected by Kerl, 3 may be quoted. The 
gases were taken from the throat of the furnaces. Table I. repre- 
sents gases from smelting thoroughly slag-roasted ore to which 
oxidized lead from the refining works and roasted matte have been 
added ; Table II., gases from resmelting the ore slag with roasted 
matte. 



TABLE I. 



TABLE II. 



Volumes. 


Ore Smelting. 


Slag- Smelting. 


i 


fN 

1 CO, 


72.72 

16.26 

10.06 

0.36 

0.59 


75.30 

17.80 

5.20 

0.10 

1.60 


75.20 

17.20 

5.40 

0.70 

1.50 


75.5 

16.6 

5.9 

2.0 


76.0 

17.4 

4.3 

1.8 
0.5 


74.8 

18.5 

3.5 

0.3 

2.6 


I 2 * 


'co 2 .: 

|CH 4 

1 H 






3 

> 

8* 


fCOo 


22.3 
13.8 


23.7 
6.9 


22.9 

7.2 


21.9 

7.8 


22.9 
5.7 


24.7 

4.7 


CO 


j z> 


c 


18.1 
29.3 


15.3 
27.1 


15.0 
26.5 


14.8 
25.6 


14.2 
25.6 


14.7 
27.0 


> g 
o 





Excess of O 


+ 2.8 


+0.6 


.... 


-0.9 


-0.9 


+0.5 


IE 

<- 





1 Balling, " MetallhUttenkunde," p. 613. 

8 Lee. cit. 

* Berg- und Hiittenmannwche Zeitung, 1880, p. 85. 



208 METALLURGY OF LEAD. 

The presence of hydrogen is explained by Schertel through the 
decomposition of marsh gas. In none of the analyses is sulphur 
dioxide taken account of. The largest amount obtained from a 
number of determinations was 0.15 per cent. The ore being slag- 
roasted very little, if any, sulphur dioxide will be set free in the 
upper parts of the furnace ; most of the little liberated in the 
lower parts is reduced to sulphur when passing through the incan- 
descent fuel and forms a sulphide. 

Schertel concludes from the analyses, in which the relation of 
carbonic acid to carbonic oxide by weight is as 4:1, that at the 
tuyeres most of the carbon is converted into carbonic acid. He 
says that atmospheric air contains, for every 100 vols, of nitrogen, 
26.5 of oxygen. If the oxygen of 126.5 vols, of air is completely 
converted to carbonic acid, there will result 126.5 vols, of gas, con- 
taining 100 vols, nitrogen and 26.5 of carbonic acid, equal to 13.25 
vols, of gaseous carbon. If on the other hand the oxygen is con- 
verted completely into carbonic oxide, the resulting gas will con- 
sist of 100 vols, nitrogen and 53 vols, of carbonic oxide, equal to 
26.5 vols, of gaseous carbon. From the tables it is evident that 
the proportion of 100 vols, nitrogen to 13.25 vols, gaseous carbon 
corresponds more closely to the figures of gaseous carbon in the 
analyses than would 100 vols, nitrogen to 26.5 vols, gaseous carbon. 
Consequently the carbon burns before the tuyeres rather to car- 
bonic acid than to carbonic oxide. 

The slags produced in the furnaces from which Schertel made 
the gas analyses contain 4.75 per cent, lime and 0.54 per cent, mag- 
nesia. This would not increase the amount of carbonic acid to 
any extent. 

The charge in passing from the throat of the furnace undergoes 
the following changes. It first loses its hygroscopic water, then 
that which is chemically combined. During the stages of incipient 
redness (525° C.) and dull redness (700° C), oxide of lead will re- 
act on sulphides of lead, iron, and silver, unless already combined 
with silica or reduced to metal. Lead sulphate may also act on 
sulphide ; its reduction to sulphide begins and that of porous fer- 
ric oxide (roasted matte). Carbonate of lead is decomposed, dolo- 
mite loses part of its carbonic acid, and the reduction of carbonic 
acid by carbon begins. 

As the charge descends through incipient cherry-redness (800° 
C.) and cherry-redness (900° C.) the reactions begun in the upper 
zone are about completed, except the reduction of carbonic acid 



SMELTING IN THE BLAST-FURNACE. £09 

and that of porous iron oxides by carbonic oxide. The action of car- 
bon on iron oxides now begins. The lead set free reacts on lead sul- 
phate, arsemate, and antimoniate, if these have not been reduced to 
sulphide, arsenide, and antimonide by means of carbon and car- 
bonic oxide. The lead also reacts on silver chloride and sulphide. 
The charge begins to become pasty ; iron and lime carbonates are 
decomposed ; any lead sulphate still existing is converted into sili- 
cate; the reduction of iron oxide by carbonic oxide decreases, that 
by carbon grows ; sulphides and arsenides rich in lead form to a 
greater extent and are in part decomposed by metallic iron. As- 
cending vapors of metallic zinc are sulphurized and oxidized. 

Passing through the stages of a clear cherry-red (1,000° C.) and 
a deep orange (1,100° C.) to a clear orange (1,200° C), fusion is per- 
fected. Ferrous oxide and lime combine with silica, setting free 
lead oxide, most of which is reduced by carbon to metal; what re- 
mains may oxidize metallic iron to magnetic oxide, sometimes 
found in matte and slag ; metallic iron acts on sulphides and ar- 
senides, setting free some lead, and forms the matte and speise, 
with prevailing iron, as tapped from the furnace. Lime and baryta 
may also act as a desulphurizing agent, carrying some calcium and 
barium sulphide into the slag ; some zinc oxide will be reduced 
and vapors ascend in the furnace. 

The melted masses separate according to their specific gravities, 
the lead collecting in the crucible and passing off through the lead- 
well; then come speise, matte, and slag, which are tapped at inter- 
vals into the slag-pot, where they separate again in horizontal 
layers. 

§61. Calculation of Charge. 

a. Introductory Remark*. — In calculating a charge for the 
lead blast-furnace, the typical slag best corresponding to the 
character of the ore i^ -elected, and the necessary amount of fluxes 
and fuel then determined. Other considerations are the amount of 
lead thai the charge will contain, the richness of the bullion to be 
produced, and the quantity of speise, matte, and slag that will 
ensue from the charge. A complete calculation will give full in- 
formation on all these points. 

To begin with the accessary amount of lead, which is expressed 
in percentage having reference to the sum of ore and fluxes: 

Charges with as little as <;.;> percent, of lead have been run suc- 
cessfully; the highest, with from 25 to 30 per cent., is hardly ever 
reached. With pure ores, containing little or no zinc, arsenic, or 



210 MET ALL URGY OF LEAD. 

antimony, and not much sulphur, it is safe to go as low as 8 per 
cent.; if these impurities are present to any extent, the charge 
should contain not less than 12 per cent. Ordinarily the lead in 
charges ranges from 12 to 18 per cent. It is to be noted that more 
lead is lost by volatilization with a charge low in lead than with 
one that is high ; the loss in silver depends mainly on the loss of 
lead and on the richness of the base bullion. 

For calculating a charge, a carbonate ore containing some 
galena may serve as an example. Its composition is : 

Si0 2 . FeO. MnO. CaO. MgO. BaO. ZnO. AL0 3 . S. As. Pb. Cu. Ag, ozs. Au, ozs. 
32.6 14.8 4.3 2.2 1.4 1.5 2.4 2.5 4.4 0.5 20.7 2.9 50.5 trace 

The typical slag shall be : 

30SiO 2 — 40FeO — 20CaO. 

The charge shall weigh 1,000 pounds and contain 10 per cent, 
of slag ; the fuel — coke — shall be 15 per cent, of the charge. The 
analysis of the iron ore shows : 

Si0 2 . FeO MnO. CaO. 

4.3 72.4 1.7 3.1; 

that of the dolomitic limestone, 

Si0 2 . FeO. CaO. MgO. 

2.7 4.5 37.3 11.9. 

The coke contains 10 per cent, of ash, consisting of 

Si0 2 . FeO. CaO. MgO. A1 2 3 . 

40.3 26.5 6.9 2.4 20.4 

Before beginning the calculation it is necessary to bring the dif- 
ferent slag-forming components of ore, fluxes, and fuel under the 
three main heads of silica, ferrous oxide, and lime. 

The atomic weights of iron and manganese being very nearly 
the same, 56 and 55, the two oxides are simply added. Both fer- 
rous oxide and metallic iron have to be considered in the calcula- 
tions: 

FeO x -^- = Fe ; Fe x -|- = FeO. 

y 7 

It will also be necessary to express the equivalents of one com- 
ponent (ferrous oxide) in terms of the other two (silica and lime). 



SMELTING IN THE BLAST-FURNACE. 211 



Let Si0 2 = c, FeO = a, CaO = b, and a-\-b-\-c = 90. 



FeO : Si0 2 : : a : c, FeO : CaO : : a : b, 

FeO = — Si0 2 . FeO = -%- CaO. 

c 6 



Under the head of lime are to be classed magnesia and baryta. 
For instance: 

CaO : MgO : : 56 : 40. 
CaO = MgO X 1.4. 

In the same way CaO = BaOx0.4. 

Some metallurgists bring also zinc oxide under the head of 
lime: 

CaO = ZnOx0.7, 

thus cutting down the lime of a slag with the increase of zinc 
oxide. 

The analyses of ore, fluxes, and coke-ash, changed as indicated, 
are: 

Si0 2 . FeO. CaO. ZnO. A1 2 3 . S. As. Pb. Cu. Ag,ozs. Au,ozs. 

Lead ore.... 32.6 19.1 10.16 2.4 2.5 4.4 0.5 20.7 2.9 50.5 trace. 

Iron ore. ... 4.3 74.1 3.10 , 

Limestone.. 2.7 4.5 53.96 

Coke-ash . . . 40.3 26.5 10.26 . . 20.4 

In figuring the charge five calculations have to be made to find: 

1. The amount of available ferrous oxide and metallic iron; 

2. The amount of metallic iron required by the arsenic to form 
Fe 5 As; 

;;. The amount of metallic iron required to combine to FeS 
with the sulphur not taken up by the copper as Cu 2 S; 

4. The amount of flux required for the 15 pounds of ash in the 
150 pounds of coke; 

5. The amount of flux required to slag the silica of the lead 
ore. 

1. Available ferrous oxide and metallic iron in the iron ore, 
100 lbs. In the Blag, 30 Si0 2 require 40 FeO. In 100 lbs. iron ore 
there are 4.3 lbs. SiO* These require 

SiO, . FeO : : 30 : 40 : : 4.3 : x; 
x = 5.7 FeO. 

The iron ore contains 74.1 per cent. FeO; deducting 5.7 gives 
68.4 available FeO, or J FeO = 53.2 available Fe. 



212 METALLURGY OF LEAD. 

2. Arsenic and iron. 100 lbs. lead ore contain 0.5 lbs. of As. 

As : Fe 5 : : 75 : 280 : : 0.5 : x; 
x = 1.86 Fe. How much iron ore is required? 
Iron ore : available Fe : : 100 : 53.2 : : y : 1.86; 
y = 3.5 lbs. iron ore. 

3. Copper, sulphur, iron. 100 lbs. of lead ore contain 2.9 lbs. 

Cu 2 : S : : 126.8 : 32 : : 2.9 : x; 
x = 0.73 S. 



Cu. 



Of the 4.4 lbs. S contained in the 100 lbs. of lead ore, 0.73 are re- 
quired for the Cu; the difference, 3.67, must be combined with Fe: 

S : Fe : : 32 : 56 : : 3.67 : y; 
y = 6.42 Fe, which corresponds to 12 lbs. of iron ore, viz: 
Iron ore : Available Fe : : 100 : 53.2 : : z : 6.42 ; 
z = 12. 

For the arsenic and sulphur of 100 pounds of lead ore 15 lbs. of 
iron ore are required. These have: 0.66 lbs. Si0 2 , 0.48 lbs. CaO, 
and 0.88 lbs. FeO. Only the non-available FeO enters the slag 
according to 30SiO 2 : 40 FeO; the rest, i.e., the available FeO, com- 
bining as Fe with the As and S to form speise and matte. The 
0.66 lbs. Si0 2 require 

Si0 2 : CaO : : 30 : 20 : : 0.66 : x; 

x = 0.44 CaO, which is balanced by the 0.48 CaO already present. 
If this were not the case, the 0.44 would have to be supplied by 
limestone: 

Limestone : CaO : : 100 : 53.96 : : y : 0.44. 

4. Coke-ash, 100 lbs. 
b. Murray's Method. — The method used resembles very closely 
the one given by Murray 1 and elaborated by Newhouse. 2 The 
analvses show: 



Desired 
Amount. 


Material. 


SiO a . 


FeO. 


CaO. 


100. 

X. 

y- 


Coke-ash 


40.3 
4.3 

2.7 


26.5 

74.1 

4.5 


10.26 

3.10 

53.96 


Iron ore 


Limestone 





1 Engineering and Mining Journal, August 13, 1887 ; March 5, 1892. 

2 School of Mines Quarterly, ix., p. 373. 



SMELTING IN THE BLAST-FURNACE. 



213 



Starting again with 100 lbs. of coke-ash, the necessary quantities 
of iron ore (x) and limestone (y) can be found by expressing the 
amounts of FeO first in terms of CaO, then in terms of Si0 2 (see 
above), and finally by putting these quantities equal to each other, 
when x and y can be easily calculated. 



FeO 



CaO. 



40 



26.5 4- 0.741a; + 0.045?/ = -^ (10.26 + 0.031a; + 0.539?/), 
x = 1.668^ — 8.80. 



FeO 



Si0 2 , 



40 



26.5 + 0.741a; + 0.045?/ = — - (40.3 + 0.043a; + 0.027?/), 
oU 

x = 39.6 - 0.001?/, 



- 8.80 = 39.6 - 0.001?/, 

y = 30 lbs. limestone, 
x = 39 lbs. iron ore. 



5. Lead ore, 100 lbs. The analyses give 



Desired 
Amount. 


Material. 


Si0 2 . 


FeO. 


CaO. 


100 
X 

y 


Lead ore 


32.6 
4.3 

2.7 


19.1 

/ 74.1 
4.5 


10.16 
3.10 
53.9 


Iron ore 


Limestone 





FeO = -£- CaO, 

19.1 + 0.741a; + 0.045?/ = ijj- (10.16 + 0.031a; + 0.539?/), 
x = 1.8 + 1.52?/. 



FeO = 



Si0 2 , 



-jo 



19.1 + 0.741a- + 0.045?y = -gj- (32.6 + 0.043a; + 0.027?/), 
x = 35.51 - 0.001?/. 



1.8 + 1.52// = 35.51 - 0.001?/, 

y = 22 lbs. limestone, 
x = 35 lbs. iron ore. 



214 METALLURGY OF LEAD. 

In summing up there are : 

15 lbs. coke-ash, requiring 
6 " iron ore. 

4 " limestone. Then the charge contains 
100 " slag. 

125 lbs. 
The difference from 1,000 lbs. =875 lbs. is to be made up by- 
ore and fluxes. Now 

100 lbs. of lead ore require to slag the 

Si0 2 , 35 lbs. iron ore -+- 22 lbs. limestone. 

100 lbs. of lead ore require to combine 

with As and S 15 

Total 172 lbs. of fluxes. 

172x = 875 ; x = 5.088, gives as charge in round figures : 

Coke-ash 15 ( = 150 lbs. of coke). 

Slag 100 

Lead ore 510 

Iron ore for Si0 2 185 

Iron ore for As and S 75 

Limestone 115 

1.000 

Figuring the pounds of each component of the charge and ad- 
ding like to like must give the slag. 

By 216.57a: = 30, the coefficient is obtained with which the totals 
of Si0 2 , FeO, and CaO have to be multiplied to obtain the desired 
figures : 30, 40, 20. The table shows that the. calculation is correct. 

By adding the different components that go to form the slag, 
665 pounds are obtained, and any change that is to be made in the 
slag must be calculated as having reference to this figure. 

For every 12.8 ounces of silver there are 102.5 pounds of lead ; 
the base bullion will therefore assay about 149 ounces silver to the 
ton. 

There are 10 per cent, of lead in the charge. 

In the charge are 2.5 pounds of arsenic, which, with 9.3 pounds of 
metallic iron, form about 12 pounds of speise. 

The 14.8 pounds of copper, requiring 3.7 pounds of sulphur, 
will form 18.5 pounds of copper matte. Deducting the 3.7 pounds 
of sulphur from the total sulphur leaves 19.7 pounds, which, with 
34.5 pounds of metallic iron, give 54.2 pounds of iron matte ; the 
total matte formed will be about 72 pounds. 



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216 METALLURG Y OF LEAD. 

The table further shows that there are 10 per cent, of slag and 
15 per cent, of fuel to the charge, thus giving all the necessary 
data. 

Before making up the charge as it goes to the blast-furnace, the 
moisture has still to be considered. If the lead ore contains, for in- 
stance, 5 per cent, of moisture, 535 pounds moist ore will have to be 
used to correspond to 510 pounds of dry ore: 

Moist Ore : Dry Ore = 100 : 95 : : x : 510 ; 
x = 537.7. 

The same is the case with fluxes and fuel. 

It is to be noted that figuring a charge according to Murray's 
formula has one great advantage over the method next to be de- 
scribed, viz., that it shows in what proportions any three classes 
of siliceous, ferruginous, and calcareous ores are best mixed so as 
become self -fluxing. 

c. Common Method. — In the common method the same ores, 
fluxes, and fuel, and the same slag as before are taken as a basis. 

The preliminary calculations, such as bringing the different 
components of ore, flux, and fuel under the heads of Si0 2 , FeO, and 
CaO, are made in the same way. The total weight (1,000 pounds) 
that the charge is to have, and with it the percentage of fuel (15 
per cent.) and slag (10 percent.) to be added are fixed. The avail- 
able FeO and Fe of the iron ore are determined as before. 

Two calculations are now necessary to determine the amounts 
of iron ore and limestone required by the coke-ash and by the 
ore. 

The analyses of the ash and the two fluxes, iron ore and lime- 
stone, are entered, as shown by the heavy type in the table below. 
The 150 pounds of coke contain 15 pounds of ash ; for these the to- 
tals of Si0 2 , Feo, and CaO are figured and entered in the table. 

There are 6.04 lbs of Si0 2 ; how much FeO is required? 
Si0 2 : FeO : : 30 : 40 : : 6.04 : x ; 
x == 8.05 lbs. FeO are necessary ; 
3.97 lbs. FeO are present. 
The difference, y = 4.08 lbs. FeO have to be added. 

To find the necessary iron ore : 

Iron Ore : Available FeO : : 100 : 68.4 : : z : 4.08, 
z = 6 iron lbs. ore. 
They are entered in the table ; their total pounds of Si0 2 , FeO, 
and CaO are figured and also entered. 



SMELTING IN THE BLAST-FURNACE. 



21: 



To the previous sum of 6.04 lbs. Si0 3 have been added, by the 
iron ore, 0.26 lbs. Si0 2 , making the total Si0 2 for which limestone 
has to be provided, 6.30 lbs. How much CaO is required ? 

Si0 2 : CaO : : 30 : 20 : : 6.30 : u, 



The difference, v 



4.20 lbs. CaO are necessary ; 

1.73 lbs. CaO are present. 

2.47 lbs. CaO, have to be added. 



To find the necessary limestone (leaving out the Si0 2 and FeO 
it contains): 

Limestone : CaO : : 100 : 53.9 : : w : 2.47, 

w = 4.5 lbs. limestone, 

which are entered with the pounds of CaO they bring to the slag. 

To see if the calculaton is correct, the pounds of FeO and CaO 
are multiplied by a coefficient (4.76, from 6.30a? — 30, x = 4.76), 
which changes the pounds of Si0 2 to 30, the percentage of Si0 2 of 
the slag aimed at. The result will be 40 FeO and 20 CaO. The 
table gives 40.08 FeO and 19.75 CaO, showing the calculation to 
give sufficiently close results. 



Material. 


Si0 2 . 


FeO. 


CaO. 


Name. 


Dry 

Weight 
lbs. 


Per 

cent. 


lbs. 


Per 

cent. 


lbs. 


Per 
cent. 


lbs. 


Coke-ash 


15 
6 
4.5 


40.30 
4.30 
2.70 


6.04 
0.26 


26.50 

74.10 

4.50 


3.97 
4.45 


10.26 

3.10 

53.90 


1.54 
0.19 
2.42 


Iron ore 


Limestone 




Total 


25.5 




6.30 




8.42 




4.15 

19.75 




Coefficient 


4.76 




29.98 




40.08 





The weights of iron ore (6 pounds) and limestone (4£ pounds) 
are practically the same as those found by using Murray's method. 

Deducting 125 pounds (the sum of coke-ash, with its iron ore, 
limestone, and slag) from the total weight of the charge of 1,000 
pounds, gives the same 875 pounds as before to be made up by the 
ore and its fluxes. 

A table, like the one below, is laid out, the analytical data are 
entered, and the calculation is made on a basis of 100 pounds of 
ore. 



218 



METALLURGY OF LEAD. 



1. The amounts of iron ore required (15 lbs.) by the As and 
S are calculated as shown on pages 211 and 212, and the results 
are entered in the table. 

2. 100 lbs. of ore contain 32.6 lbs. Si0 2 , for which the necessary 
iron has to be provided : 

Si0 2 : FeO : : 30 : 40 : : 32.6 : x ; 
x = 43.46 lbs. FeO are necessary ; 
19.10 lbs. FeO are present. 
The difference, y = 24.36 lbs. FeO have to be added. 

To find the necessary iron ore : 

Iron ore : Available FeO : : 100 : 68.4 : : z : 24.36 ; 
z == 35 lbs. iron ore. 

3. To the 32.60 lbs. Si0 2 of the ore, have been added from 
the two additions of iron ore 0.66+1.50 = 2.16 lbs. Si0 2 , making 
the total of 34.76 lbs. of Si0 2 , for which lime has to be provided : 

Si0 2 : CaO : : 30 : 20 : : 34.76 : u; 
u = 23.17 lbs, CaO are necessary ; 
11.17 lbs. CaO are present. 
The difference, v = 11.45 lbs. CaO have to be added. 

To find the necessary limestone (leaving out the Si0 2 and FeO 
it contains): 

Limestone : CaO : : 100 : 53.90 : : w : 11.45; 
w = 21 lbs. limestone, 
which are entered upon the table. 

Adding the pound-columns of Si0 2 , FeO, and CaO, and multi- 
plying by 0.86, proves that the calculation is correct. 



Material. 


Si0 2 . 


FeO. 


CaO. 


Name. 


Dry 

Weight 
lbs. 


Per 

cent. 


lbs. 


Per 
cent. 


lbs. 


Per 
cent. 


lbs. 


Ore 


100 
15 
35 
21 


32.60 
4.30 
4.30 
2.70 


32.60 
0.66 
1.50 


19.10 

74.10 

74.10 

4.50 


19.10 

0.88 

25.93 


10.16 
3.10 
3.10 

53.90 


10.16 
0.48 
1.08 

11.32 


Iron ore for As and S . . 

Iron ore for Si0 2 

Limestone 




Total 


171 




34.76 




45.91 


23.04 
19.81 




Coefficient 


0.86 




29.89 




39.58 











SMELTING IN THE BLAST-FURNACE. 219 

The figure found for iron ore (35 lbs.) is the same as with 
Murray's formula; that for limestone is slightly lower (21 vs. 22 
lbs.), as the Si0 2 and FeO contained in the limestone are left out, 
making the available CaO 53.90, which is slightly too high. 

If the items of the ore charge are now multiplied by 5.088 (as 
on page 217) and those of the coke charge added, the sum of 1,000 
pounds, the entire charge, will again be obtained. 

GENERAL SMELTING OPERATIONS. 

§ 62. Blowing-in. — This consists of three parts: Warming 
the crucible, filling the furnace, and starting the smelting. 

The warming of the crucible in a new furnace must be done 
slowly and with great care, so as to raise the temperature gradual- 
ly; otherwise the quickly escaping moisture will crack the masonry 
and allow the lead in the crucible to percolate. The practice of 
metallurgists differs in the manner of warming the crucible and the 
amount of time given to it. The writer has, whenever it was pos- 
sible, taken forty-eight hours to warm the crucible, and has been ac- 
customed to proceed in the following way: the water is turned in- 
to the jackets so as to fill them and have just a little overflow; then 
the flue leading to the dust chambers is closed, and the damper on 
top of the stationary stack raised, or the movable stack put in 
place, so that the gases may pass off into the open air. 

With some of the latest furnaces the iron plate for closing the 
furnace is lowered, and the joint covered with fine ore, the gases 
formed, when blowing-in, being allowed to pass off into the dust 
chamber. This can only be done if the furnace is blown in with 
very little or no charcoal, which means having the crucible full of 
molten lead before the charge is introduced. If the furnace were 
filled to the top of the jackets with charcoal before adding the 
blowing-in charges, the gases, when the blast is let on, being rich 
in carbonic oxide, would be liable to ignite and damage the flue 
and duet chamber. 

A wood fire is made in the bottom of the furnace that will not 
reach half-way up the crucible. If it is kept going for a few hours, 
always replenishing the wood, ashes will have collected in the 
crucible. These, being bad conductors of heat, have to be re- 
moved in order that the burning wood may be in contact with the 
bottom. When raked out by means of a hoe from the breast of 
the furnace, a new fire is kindled. After from three to four hours 



220 METALLURGY OF LEAD. 

too many ashes will have accumulated in the furnace for the heat 
to have the desired effect, and the crucible is cleaned out again. 
While the crucible is being dried and warmed, the lead-well is 
filled with glowing charcoal, and the basin itself covered by a piece 
of sheet iron, so as to admit only a little air, thus preventing the 
charcoal from being burned quickly. Similarly the breast of the 
furnace is closed with loosely set bricks, by which the draught is- 
checked and too quick combustion of fuel on the surface prevented. 
The heating is continued for twenty-four hours, when the outside 
of the crucible will feel warm to the touch. This shows that all 
the moisture is expelled and that the crucible can stand a high 
heat without endangering the brick-work. 

The second warming has for its object the heating of the cru- 
cible to a good red-heat, and requires carbonized fuel and blast. A 
new fire is kindled and some charcoal added ; the blower is started, 
and all the tuyere-bags are tied up or wound up, or the blast-gates 
are closed, except that of the tuyere nearest the breast to be used. 
This is connected with an iron pipe inserted into the crucible from 
the breast, and the blast allowed to play on the charcoal till it is 
well ablaze. A second layer of charcoal is added and, when it is 
fully aglow, a third one and so on till the crucible is filled to the 
jackets. The iron pipe is then inserted deep into the glowing coal,, 
in order that the blast may reach the bottom of the crucible. Mean- 
while the furnace-man works at intervals with an iron bar and 
a hoe, turning the coal over and moving it from front to back and 
vice versa, so as to get it all into a perfect glow. When this is 
accomplished, the pipe, of which often a small part has been melted 
off, is withdrawn, and the furnace is let alone for an hour or two. 
The pipe is again introduced and the charcoal burned down, the 
furnace-man stirring it to bring all parts into contact with the 
blast. The ashes are removed, and the crucible and well thoroughly 
cleaned. The heating is repeated from three to four times in 
twenty-four hours. The outside of the crucible will then have be- 
come too hot to be touched with the hand. 

The second step is the filling of the furnace. This must be done 
with care, in order that a crust may not form later on the lead, 
when the blast has been started, as it is very troublesome to re- 
move, and sometimes remains in the furnace during an entire run. 
To avoid a crust, it is essential to have a clean crucible entirely 
filled with red-hot lead. If half filled with a mixture of charcoal- 
ashes and small bits of charcoal, a dead layer will be formed be- 



SMELTING IN THE BLAST-FURNACE. 221 

tween the lead and the slag which will soon harden to a crust and 
attract small obstacles that would otherwise be carried out by the 
slag. The proper mode of filling a blast-furnace depends a good 
deal on the kind of fuel, and is to some extent individual with the 
metallurgist. After supplying the crucible with the necessary 
lead, which is either charged from the top or melted down from the 
breast of the furnace, the first charges will consist of an easy- 
smelting slag with much fuel and the necessary fluxes. Ore-charge 
gradually replaces the slag-charge, and the high percentage of fuel 
will be cut down until finally the normal charge is reached. It is 
better for a blowing-in slag to be glassy than crystalline, as it 
melts easier, and a slag consisting of 30 Si0 2 , 40 FeO, and 20 CaO 
is probably the best. 

Two methods of blowing-in may be mentioned ; they both re- 
quire charcoal. 

For the first, the thoroughly heated crucible is gradually filled 
with charcoal, the blast playing on it till it is all aglow ; then the 
breast of the furnace is put in, the tuyere openings are closed, and 
the rest of the charcoal is added from the feed-floor, so as to reach 
about one foot above the tuyeres. Then follows a bed of coke 
one foot thick, which will reach about to the top of the jackets. 
Upon this are charged lead and coke, with some slag and the 
necessary fluxes. When all the lead has been given that is required 
to fill the crucible, the furnace is filled with alternate layers of 
fuel and half ore-charge and slag-charge, using the same amount 
of fuel for ore as for slag. 

The necessary amount of iron ore and limestone are calculated 
in the same way as in making up the regular blast-furnace charge. 

The tuyere-holes in the jackets are now cleared, the lead-well is 
partly uncovered, the tuyere-pipes are put in place, and the blower 
is started slowly, the number of revolutions as well as the pressure 
of the blast being noted. 

Too much emphasis cannot be laid on starting the blast slowly. 
If the blower is allowed to make too many revolutions per minute, 
the result is that the fuel burns quickly, the heat rises, and is not 
concentrated at the tuyere-level, and the gases rush upward and 
do not thoroughly warm the charge ; the pressure of the blast. 
at the same time lowers the level of the lead in the crucible, with 
the charcoal ashes floating on top. When finally slag and matte 
come down, they become partially chilled, because they sink too 
much below the level of the tuyeres, and thus the dreaded blowing- 



222 METALLURGY OF LEAD. 

in crust is formed. For this reason the blower should always be 
started up slowly, and the pressure not allowed to be much over 
one ounce until slag appears before the tuyeres. The normal 
pressure of the blast is not reached until seven or eight hours after 
the blower is started. 

When the blast has been turned on a flame issues from the 
siphon-tap till the lower aperture is closed by the lead. Care 
must be taken that the passage of the well does not become 
choked with ashes, which are carried out by the flame in con- 
siderable quantities. The lead rises slowly as the bullion melts 
down in the furnace, and while often at first a little cool, it soon 
becomes hotter, and, when it reaches the top of the well, ought to 
bebright-red. Shortly after the blast has been let on, the flow of 
water into the jackets must be increased in order that it may not 
boil. Soon, however, when the slag charged begins to melt and 
come down, the jackets become coated and thus cooler. The large 
excess of water required just before is turned off, and less than the 
normal amount will be sufficient for an hour or two. On the feed- 
iloor the charges will at first sink quickly when the blast has been 
turned on, and it may be sometimes advisable to continue feeding 
alternately ore-charge and slag-charge until the furnace has been 
entirely filled with this mixture, and only then to begin with nor- 
mal ore-charge. This is, however, exceptional. The smoke issu- 
ing at first from the top of the charge will be black, and gradually 
become lighter, and decrease in amount. The furnace-man watches 
through the peep-hole in the tuyere-pipe to see when it is time to 
tap the slag. This will be indicated by the appearance of a little 
"blue flame of burning carbonic oxide gas. 

The ashes of charcoal have now to be removed from the furnace, 
the lead having lifted them up from the bottom to about the level 
of the tap-hole. They float on the lead, and are liable to prevent 
the necessary contact of lead and slag or lead and matte in the 
furnace. Generally the slag carries out the ashes, but sometimes 
this has to be assisted, which is done by entering with a rod 
through the tap-hole and loosening the ashes where they are in- 
clined to accumulate and adhere, thus stirring them into the slag. 
It is often advisable after stirring with a rod to let the blast blow 
through the tap-hole, as it will blow out a good many ashes. If, 
nevertheless, a crust of ashes and slag should form on top of the 
lead, it is necessary to break it by thrusting bars through it and 
lifting it up while it is still soft. If it hardens, the furnace will 



SMELTING IN THE BLAST-FURNACE. 223 

have to be stopped, and the breast taken out before the crust can 
be removed (see § 66). The indications of a crust are that the 
lead in the well becomes dark and does not play freely with the 
blast. The first is easily seen ; the second can be discovered by 
removing the plug or cap of the peep-hole in the tuyere-pipe just 
above the lead-well. If the lead plays freely, it will sink in the 
well; if not, it will remain immovable. By inserting a rod through 
the tap-hole the position of the crust can be felt. 

If everything goes well, the furnace will have its full blast in 
about seven hours after starting, the slag its normal heat next day. 

One of the main points in filling a furnace is to be quick about 
it ; therefore as many men are put to work as is possible. A fur- 
nace 42 by 108 inches can be filled in from 30 to 40 minutes by 
two sets of feeders, of four each, helped by as many wheelers as 
may be required to bring the necessary materials to the furnace. 
If there is any delay in charging, the crucible may cool and a crust 
form. For this reason no weighing is done at the furnace, but the 
charge is measured by the number of scoopfuls and shovelfuls, the 
average weight of one having first been determined by weighing 10 
scoopfuls of coke and. 10 shovelfuls of iron ore and limestone. 

With good charcoal this method of blowing in is quite satis- 
factory ; a crust very rarely forms. With charcoal of an inferior 
grade the second method is used. It is as follows : when the 
crucible has been well heated up and cleaned out as described, the 
breast is put in, the tuyeres in the jackets are closed, and the lead- 
well is covered with sheet iron, its lower side being coated with a 
clay lute and the upper one weighted with a couple of bars of lead. 
Charcoal is fed from the feed-floor until it reaches the top of the 
jackets, and covered by a bed of coke about one foot in thickness. 
A man descends into the furnace and spreads out evenly the suc- 
cessive charges of bullion, coke, slag, iron ore, and limestone. 
Then follow the regular ore charges. When the furnace has been 
filled, which takes about two hours, the charcoal is kindled from 
the tuyeres and a gentle blast started. The work now proceeds in 
the usual way. When the lead, coming down in the furnace, has 
closed the lower opening of the well, the cover is removed. 
Whether this has occurred can be found by removing from time 
to time the iron rod which closes the tap-hole of the lead-well. 
The first slag is tapped about 25 minutes from the time that the 
lower opening of the well is closed by the lead, which shows how 
quickly the crucible fills. 



224 METALLURGY OF LEAD. 

The reason for the good results attained by this method may be 
explained in the following way : as the entire lower part of the fur- 
nace is closed during the filling, and the charcoal is not kindled un- 
til just before the blast is started, it cannot burn in the crucible 
and form the dreaded mixture of ash and fine coal. At the 
tuyeres much heat will be generated by the combustion of the 
charcoal. A slight effect may be communicated downward, but it 
proceeds principally upward. The first lead that melts and trickles 
down is very hot when at the tuyere-level, but cools somewhat be- 
low it. As more follows, the first becomes heated again, so that 
finally the bottom of the crucible is filled w T ith liquid lead, 
although it is not yet very hot. The unburn ed charcoal is raised 
gradually towards the tuyeres, where it burns completely, leaving 
no mixture of ash and fine coal, and gives a very high temperature, 
which prevents the forming of a crust. The lead as it rises in the 
crucible becomes always hotter from the continual addition of red- 
hot lead from above and its approach to the tuyere level, as well as 
by its contact with the hot slag. 

Both of these methods require passably good charcoal, if not 
the best. If this cannot be had, it is replaced by split wood which 
is thoroughly dry and cut into pieces of suitable length. "When 
w r ood is used instead of charcoal, it is burned down in such a way 
as to make fresh charcoal in the crucible. Thus, when this is 
sufficiently heated and cleaned out, the last fire is made and the 
wood burned down quickly with the blast only so far as to be well 
charred, fresh wood being added and the burning down continued 
until the crucible is filled with charred sticks. The breast is put 
in place and a two-foot bed of coke given. Upon this come the 
regular blowing-in charges, the double amount of coke requiring 
an extra quantity of slag, iron ore, and limestone. 

If neither charcoal nor suitable sticks of dry wood are available, 
as is often the case in Mexico, it will be necessary to melt down 
the lead in the crucible before filling the furnace, as filling the 
crucible with coke is nearly sure to form a very disagreeable crust, 
which it is difficult to remove afterwards from the breast. If neg- 
lected, it grows rapidly, and the result may be a crucible frozen 
up solid. The mode of procedure is to fill the well, after it and 
the crucible have been heated and cleaned with glowing coals, and 
to pile them high up over it, so that it shall keep hot. Then as 
many bars of lead as the crucible will hold are introduced from the 
breast by sliding them in on a board. A fire is kindled on top of 



SMELTING IN THE BLAST-FURNACE. 225 

the lead ; when that is melted down, the floating ashes are removed,, 
a new fire is made, more lead melted down, and this repeated untiL 
the crucible is full of lead. The last ashes are raked out, the breast 
is put in, and enough of the best brushwood that has been sorted 
out is charged from the feed-floor to reach well above the jackets. 
On top of this comes a two-foot bed of coke, to which eight or ten 
bars of lead are added, in order that this, coming down hot, may 
help to heat the lead in the crucible. Then come the usual slag 
charges and slag-ore charges. 

The bullion which used generally to be charged from the top is 
now more commonly introduced from the bottom. The lead is 
melted down in the warmed crucible, as just described, the breast 
put in place, and charcoal fed to the top of the jackets ; then the 
coke-bed is given and the furnace filled, with the addition of a 
few bars of lead with the charge. 

Some metallurgists avoid charcoal entirely. The crucible is 
warmed with wood, the lead melted down with coke, all the 
clinkers are removed, a good coke fire is started on the lead, and 
the furnace filled with coke to the top of the jackets. Then a few 
slag charges are given and the furnace is filled with an easy ore 
charge running high in lead. This is soon changed for the regular 
charge. 

The writer has never blown in a furnace without using lead. 
Halm l says that, when no lead can be had, he closes the syphon- 
tap where it enters the crucible with a clay plug, removing it when 
the lead begins to flow out with the slag. He adds that bottom 
crusts are unavoidable. When sufficient lead has been produced 
to fill the crucible, the furnace is blown out and started up again 
in the normal way. 

A method of blowing-in described by Henrich 2 refers rather 
to a furnace where metal is tapped from the bottom than to one 
having a lead-well. It is given here because it describes the 
blowing-in of ;i blast-furnace used for concentrating matte, which 
is tapped from the bottom. The crucible is warmed, filled with 
charcoal up to the level of the tuyeres, more fuel is given, and 
then blowing-in charges are fed, followed by ore charges. When 
the furnace is filled to from one-half to two-thirds the distance 
between the tuyeres and the charging floor, the front is closed. 

1 "Mineral Resources of the United States," United States Geological 
Survey, 1882, p. 342. 

8 Engineering and Mining Journal, April 11, 1885. 



226 



METALLURGY OF LEAD. 



All the tuyeres excepting one or two (according to the size of the 
furnace) opposite the metal tap (2 inches wide) are closed, the pipes 
inserted, and the blower is started. The flame passes through the 
tap-hole, blows out ashes, and heats the bottom. The tap is kept 
open by inserting a rod until slag begins to flow, when it is 
closed with a clay plug. The other tuyere pipes are now intro- 



TABLE OF BLOWING IN CHARGES. 



Charges, 
number. 


Coke, c 
scoops. g( 


/har- 
5oal, 
oops. 


Iron ore, 
shovels. 


Lime- 
stone, 
shovels. 


Bars of 
bullion, 
number. 


Slag, 
shovels. 


Size of 
furnace, 
inches. 


Crucible 
filled 
from 


1 


20 




3 


2 


6 








3 


15 




3 


2 


6 


1 






5 
5 

2 


10 
10 
10 




3 
3 
3 


2 
2 
2 


8 
7 
5 


2 
2 
5 


36 in 
diam. 


above. 


5 


10 




half ore-charge. 


3 


6 






1 


10 




half ore-charge. 


.... 


6 












regular ore-charge. 






1 


10 








6 








3 


8 




Vz 


K 


6 


1 






4 


7 


. . . 


% 


X 


6 


2 






4 


6 


1 


% 


Va 


6 


2 


36X84 


above. 


2 


5 


2 


X A 


X 


5 


2 






2 


5 


2 


half ore-charge. 


2 






4 


4 


3 


half ore-charge. 


2 












regular ore-charge. 






1 


20 




3 2 6 








3 


15 




3 


2 


6 


1 






10 


15 




3 


2 


6 


2 


36x84 


above. 


5 


12 




3 


2 


3 






4 


6 




half ore-charge. 


4 












regular ore-charge. 






1 
12 


25 I . 

12 1 


6" 


1 11 f 15 I 

1 1 1 1 15 I 
regular ore-rharge. 


12 

12 


42xl20| above. 


1 
15 


25 1 
9 1 


... 


11 1 1 20 

1 1 1 1 2 


■m 


42xl20| below. 








regular ore-charge. 







duced at short intervals one after the other. Once or twice slag 
is tapped from the metal tap ; it is then allowed to accumulate 
in the crucible and tapping is begun at the slag tap. In this way 
the crucible becomes thoroughly heated. 

The only instance coming under the writers observation of 
slag being taken out from the lead-well was at a desilverizing 



SMELTING IN THE BLAST-FURNACE. 227 

plant, when very rich bullion, obtained by smelting zinc-crusts in 
a circular 36-inch furnace (§ 112), had to be exchanged quickly for 
low-grade bullion. The furnace was fed with a number of slag 
charges to half fill it ; then sufficient low-grade litharge and slag 
was added to fill the crucible again with lead. When the slag 
charges had about reached the tuyeres and much slag was being 
formed, the lead was all dipped out of the well and then the slag, 
until finally the lead of the litharge appeared and filled the well, 
when slag was tapped again in the usual way. The rich bullion 
was very little diluted and could go to the cupelling furnaces. 

The making up of blowing-in charges varies a great deal. The 
preceding table gives a few examples. A scoop of coke weighs 
12 pounds, one of charcoal 4, a shovel of iron ore 15 pounds, of slag 
12 pounds, and of limestone 10 pounds. 

§ 63. Regular Work on the Charging Floor. — The work on 

the charging floor consists in bringing ore, flux, and fuel from the 
bins to the scales, weighing out the required amounts, dumping 
them, and feeding into the furnace. 

Ores and fluxes are brought to the scales either in trucks or 
wheelbarrows ; the former are filled through chutes from the bins, 
the latter by shovelling ; coke and charcoal are nearly always 
brought in wheelbarrows. 

Everything that goes in the furnace must be accurately weighed. 
No thorough work can be done without. An approximate weight 
can be obtained by taking ten or twenty shovelfuls and reducing 
this weight to that of one, and then making up the charge by 
measure. But shovelfuls differ according to the men who make 
them, and the same man will not always shovel a uniform amount — 
say, at the beginning and toward the end of his shift. With the 
shovel system the furnace is always liable to get out of order, and 
the metallurgist has no means of accurately determining the cause. 
As seen from § 62 the shovel system is used in ore smelting in 
blowing in a furnace, where the filling with slag charges must be 
done as quickly as possible ; as soon as ore is substituted it must 
be regularly weighed. In refining works, where certain lots of lead 
by-products arc treated in a blast-furnace with slag alone, the 
shovel system is often round. This is permissible on account of 
the small amount of new slag that is being formed in the blast- 
furnace, but even then weighing proves more satisfactory. 

The scales in use have a capacity of from 1,000 to 1,500 pounds. 
They are of two kinds : the single-beam scale, where the weights 



228 



METALLURGY OF LEAD. 



have to be shifted for every weight taken, and the multiple-beam, 
where by pulling a lever on the outside of the case the one of 
the four, six, or eight beams which bears the desired weight can 
be set free. By the multiple-beam errors are avoided, as the fore- 
man sets the weights and locks the box surrounding them, the 
workman having only to pull the levers in their regular order. 

In making up a charge and in distributing it in front of the 
feed door the fuel is always kept and fed separately, while ore, 
fluxes, and slag are sometimes mixed and sometimes kept separate. 
With small furnaces — say, 36 by 60 inches — ore, fluxes, and slag are 

Tramway 



Two charges 
charcoal. 


Two charges 
coke. 


(Half charge 
I ore mixture. 




Half charge ) 
ore mixture.) 


( Charge of 
( .fluxes. 


Furnace-mouth. 


Charge of | 
fluxes, i 


j Half charge 
( ore mixture. 




Half charge 1 
ore mixture, S 



DISTRIBUTION OF CHARGE ON FEED-FLOOR. 

usually dumped into one heap on the long side of the furnace, and 
the fuel has its place on the short side. With a large furnace (36 
by 108 or 33 by 120 inches) ore and fluxes are sometimes distributed 
in separate heaps on the long sides of the furnace. lies, 1 for in- 
stance, adopted the following disposition of the charge on the fur- 
nace floor (Figure 145). The numbers denote the order in 
which the materials are supplied to the furnace. The feeder first 



Engineering and Mining Journal, March 24, 1883. 



SMELTING IN THE BLAST-FURNACE. 229 

charges one-half of the charcoal pile and on top of this the ore 
mixture and the fluxes 1, 3, and 5 on the same side of the furnace; 
then he gives one-half of the coke pile on the opposite side, which 
is followed by the rest of the ore mixture and the fluxes 2, 4, 
and 6. 

The mode of distributing the charge in the furnace deserves 
great attention, as by wrong feeding a good charge can be spoiled 
and a furnace be put out of order. Fuel and ore charge go into the 
furnace in alternating layers, beginning always with the former. 
The fuel must never be exposed to the air, but must be well cov- 
ered by the ore charge. In feeding, the fuel, especially charcoal, 
should be kept near the centre, and the charge distributed more 
towards the walls of the furnace, and, finally, the finer parts fed 
nearer the walls than the coarser ones. The reason for this is that 
in the descent of the charge the lighter fuel is liable to be pressed 
towards the walls of the furnace; further, as there is more friction 
between charge and furnace wall than between the parts of the 
charge itself, it will sink quicker in the middle while descending, 
-and pack at the sides. If ascending gases are to pass evenly 
through the charge, their passage must be assisted in the centre, 
and retarded at the sides. Lastly, with carbonate ores an intimate 
contact of readily reducible lead oxide with the fuel in the upper 
part of the furnace must also be avoided, as it would assist in re- 
ducing it to metallic lead, and thus increase the loss by volatiliza- 
tion. Guyard 1 thinks that this distribution of fuel towards the 
centre and of charge towards the furnace walls favors the growth 
of accretions immediately above the jackets, and advises charging 
the fuel alternately towards the centre and the sides. He says the 
effect of this would be that the accretions would begin to form 
higher up in the furnace and could be removed more easily; but 
Mould not the top of the furnace easily become hot? In feeding a 
charge that contains much fine ore (sand carbonates) special pre- 
cautions are taken to prevent this from trickling through the charge, 
as it fills up the pores of the coke, and, arriving in a crude state at 
the smelting zone, chills the furnace and often fills the tuyeres. 
Glenn 2 describes what he calls a filter charge. It is not uncommon 
with lead furnaces, and consists in making the charge large, and in 
feeding gradually on the heavy coke bed first the coarse parts, then 

1 Emmons, "Geology and Mining Industry of Leadville," United 
States Geological Survey, monograph xii., p. 665. 

2 Engineering and Mining Journal, July 19, 1884. 



230 METALLURGY OF LEAD. 

finer ones, and last of all the fines here and there, as the charge 
settles in the furnace. The finer the ore, the more uniform in size 
should the slag and fluxes be to make a good filter. Fine ores are 
often fed where the fumes are strongest, thus helping to equalize 
them. The distribution of ore and fluxes in a charge may effect 
its fusibility. Henrich l gives some very interesting experiences 
on this point, and calls attention to the fact that the formation of 
the slag, while the charge is descending in the furnace, is a gradual 
process. Pieces of ore and flux containing the necessary constitu- 
tents to form slag of a low-melting point may do this in the upper 
parts of the furnace if in close contact with one another. The 
liquid slag will then eliquate and take with it from the hotter zone 
below less fusible parts of the charge, thus forming prematurely 
the intended slag. If in charging, the ore is followed by a flux of 
similar composition, very liquid slag cannot form in the upper 
parts of the furnace, but slag requiring a higher temperature for 
its formation will be formed lower down, and will combine with 
the liquid slag at the tuyeres and form the normal slag. Thus, for 
instance, a siliceous ore should be followed by a siliceous iron flux, 
and the dissimilar flux — limestone — come last. In this way a higher 
temperature will prevail in the smelting zone than would if ore and 
fluxes were mixed and fed together into the furnace. Henrich ex- 
plains the good effect large charges often have in furnaces as com- 
pared to small ones, by the separate charging of ore and fluxes Avith 
the former, while with the latter, ore and fluxes are usually dumped 
in one heap in front of the feed door and charged together into 
the furnace. 

A furnace is in a good condition on top when this is cool, when 
the charges sink regularly and evenly, and the fumes ascend uni- 
formly, or a little more densely at the sides than at the centre. 

The labor required on the feed-floor varies with the size and 
number of the furnaces and the conveniences for handling materials. 
A furnace from 33 to 42 inches between the tuyeres and from 84 to 
120 inches in length requires per shift one feeder and two wheelers ; 
with ten furnaces under the same conditions are required ten 
feeders, but only sixteen wheelers. Both generally work twelve- 
hour shifts, although sometimes the feeders work only eight. 

The tools required by the feeder are a square-pointed, long- 
handle shovel, and a scoop or fork for the fuel, a six-pound napping 

1 Engineering and Mining Journal, December 27, 1890 ; June 6, 1891. 



SMELTING IN THE BLAST-FURNACE. 23 1 

hammer for breaking up the coke, and a stiff broom, several heavy 
1^-inch steel bars of different lengths, the longest ones reaching 
from the top of the jackets up to 4 feet above the feed floor, and 
several ten-pound double-faced sledges. The bars and sledges are 
used in cutting out wall accretions (§ 66), and one set of tools is 
sufficient for a number of furnaces. The wheelers require several 
iron wheelbarrows for ore and fluxes and a wooden barrow or buggy 
for fuel, several round-pointed shovels, a scoop or fork,"several coarse 
brooms, and generally a pick. A definite number for most of the 
tools cannot be given, as it varies too much with the general ar- 
rangement of the feed-floor. 

§ 64. Regular Work on the Furnace Floor.— This consists 

in regulating the water supply, taking care of the tuyeres, tapping 
speise, matte, and slag together into the slag-pot and wheeling it to 
the dump, tapping the lead either first into the lead-pot or ladling 
it directly into the moulds, and piling the bars of base bullion upon 
the ground or on a truck. 

The water of the jackets is kept at about 70° C, the usual test 
being that the hand can be quickly passed through the outflowing 
water without being scalded. Any irregularity in the temperature 
of the jackets indicates that the coating on the inside is thicker in 
some than in others, and thus that the smelting proceeds unevenly ; 
the slower descent of the charges on the cold side will corroborate 
this on the feed floor. 

The pressure of the blast is constantly watched and all the 
changes in the gauge noted. It is regulated by means of a damper. 
If the furnace has its own blower, its revolutions are also counted. 

The appearance of the tuyeres forms a good indication of the 
inner condition of the furnace. It is not necessary that the entire 
mouth of the tuyere should be bright ; it is often covered by a 
lhin scale of slag, showing a star-like brightness in different places. 
If it becomes quite dark, a rod is inserted and the slag pierced to 
see if it has grown too thick. In this case repeated poking only 
aggravates matters, as a "nose "of chilled slag forms, reaching 
into the furnace. (For correction of this evil see § 66.) 

The tapping of the slag is done with a square-pointed steel bar. 
The slag runs over the slag-spout into the slag-pot; when full, 
the tap-hole is stopped with the stopping-rod. As regards liquidity, 
matte conies first, then speise, and last the slag. When flowing 
from the spout they can be readily distinguished. The clay stop- 
per in the tap-hole ought to he sufficiently soft to be perforated by 



232 METALLURGY OF LEAD. 

pricking it with the bar, but oftener it requires a few taps or strokes 
with a sledge. With a hard tap it is sometimes advisable to insert 
a small piece of charcoal into the clay stopper so that it protrudes 
slightly. On shutting off the slag, the charcoal enters the opening 
made and is followed by the clay. On tapping again very little 
sledging will be necessary. The use of this charcoal is kept up 
until the correction in the charge or the feeding has softened the 
tap. The taking of samples of slag will be discussed in § 81. It is 
a rule always to keep some slag in the furnace. At certain inter- 
vals, however, the lead and slag are allowed to rise until the little 
blue flame, seen through the tuyeres, indicates that the slag has 
nearly reached that level, when it is all tapped out to see how many 
pots of slag the furnace holds. If there is less slag than usual, it 
shows that the region of the tuyeres is not as free as it ought to be. 
It is generally desirable to accumulate the speise and matte in as 
few slag-pots as possible, as, when they run out with the slag, they 
are splashed, especially the matte, against the cool sides of the pot. 
The separation of matte and slag being complete only near the 
centre of the pot, where they remain liquid for some time, an 
outer ring of rich slag, assaying often 15 ounces and more to the 
ton, will be obtained when the bulk of the slag is poor. This 
occurs especially with acid slags ; with slags that are very fluid 
the difference in outer and inner slag is not so great. If the furnace 
has a clay breast, the slag is tapped when it has accumulated in 
the furnace a little below the ordinary tap-hole. All the speise 
and matte, and perhaps a little lead, will run out and be collected in 
one pot, and all the slag of this pot will be saved. After a number 
of tappings from the ordinary tap-hole, regulated by the speise and 
matte the furnace is making, the tapping lower down is repeated. 
If the furnace has a water-cooled cinder-notch (tapping-jacket), 
the level to which the slag is allowed to rise is regulated by the 
lead alone, as a jacket with two tap-holes is not much used. 

The loss of metal by imperfect separation of matte and slag is 
completely avoided by the use of an overflow-pot (§ 49, h) or a 
catch-pot having a tap-hole (§ 49, h). 

It is of importance for the good running of a furnace that the 
level of the lead should %e kept high. If too low, the speise and 
matte will be too tar removed from the zone of fusion, and thus 
liable to chill and form a crust. The less lead there is in the charge 
the higher must the level of the lead in the crucible be kept, that 
it may not cool, as it is not frequently replaced by fresh lead that 



SMELTING IN THE BLAST-FURNACE. 233 

has just passed the hottest part of the furnace. With a charge con- 
taining from 7 to 9 per cent, of lead, the lead in the well is kept often 
as high as 5 inches above the slag-tap. It used formerly to be taken 
out altogether by dipping, but if this is not done with judgment, 
too much is apt to be removed, which causes trouble. The- amount 
is to be regulated by the number of charges that have been fed into 
the furnace. In ladling the hot lead from the basin of the well 
into moulds, most of the dross held in solution rises to the surface 
when the bar cools, and is removed by skimming, but some always 
remains. This dross often causes dispute between buyer and 
seller as to the silver contents of a given shipment of bullion. 
The dross assays much lower in silver than the clean lead, and as 
it is not uniformly distributed in the bar, the sampling becomes ir- 
regular (§§ 71 and 72). Then it is the pride of the furnace-keeper 
to produce bars with a clean surface. This has led to pouring the 
mould one-quarter full of lead, then adding dross, and covering it 
up with a layer of clean lead. The trouble can be removed by al- 
lowing the lead to accumulate in the well, which is on purpose 
built up higher than the top of the crucible, and tapping it at inter- 
vals into a cast-iron pot. The lead not being dipped out, cannot 
sink lower than the level of the tapping-hole, which is on a line 
with the wall of the crucible, and can be kept as high as desired. 
In the cast-iron pot the dross is skimmed off, and the cooled lead is 
then ladled into a series of eight or ten moulds placed in a row 
alongside the furnace on two rails, supported by two wooden horses. 
Nevertheless, many metallurgists prefer to dip the bullion from 
the lead-well. The reason is that, if the lead is tapped at long inter- 
vals from the well, a large number of bars, say ten, will be taken out 
of the furnace at once, and the charge will sink suddenly to occupy 
the space set free by the lead, w r ith the result that particles of un- 
melted charge pass too quickly through the melting-zone to be fused, 
and often form the nucleus for a crust. They therefore dip the 
lead from the well as fast as it forms; i.e. y the dipping is regulated 
by the number of charges fed into the furnace and the percentage 
of lead it contains. Dipping and tapping the lead are both used; 
the dipping may perhaps be preferable where charges run low in 
lead, say 8 per cent. Another reason for dipping the lead is that 
the refiner pays for the dross as lead. If the smelter skims the 
dross, he not only is not paid for it, but has to work it into copper- 
matte, and copper-refiners pay nothing for lead and only 93 per 
cent, for the silver of the matte. This is probably the main 
reason why many smelters prefer dipping to tapping. 



234 METALLURGY OF LEAD. 

When the bars have solidified they are marked with the running 
lot-number, removed from the moulds with a pick, and piled up to 
be sampled (§ 72), weighed, and shipped. 

A furnace is in good working order below when the tempera- 
ture of the jackets is uniformly high, the pressure of the blast 
does not fluctuate, and the tuyeres remain bright, having only short 
"noses." The furnace should within a given time produce always 
the same number of pots of slag, the tap-hole be neither too hard 
nor too soft, and the lead in the well be of a bright-red color, 
play with the blast, and sink slightly every time a pot of slag is 
tapped. 

On the furnace floor are required for every furnace (from 33 to 
42 by from 84 to 120 inches) one furnace-keeper, one tapper who 
looks after the lead, and from one and one-half to three pot -pullers, 
according to the size of the dump and the manner of disposing of 
the slag. 

The tools required by the furnace-man are 2 tapping-bars (6 or 
8 feet long, of f-inch steel), 2 stopping-rods (10 feet long, of -J-inch 
iron, with a disc 2^ inches in diameter for the clay plug), 2 iron 
rods (12 feet long, of f-inch iron) for special cases, 3 or 4 heavy (1— 
1^-inch) steel bars about 6 feet long, and 2 ten-pound double- 
faced sledges ; a 2-inch steel bar 8 feet long is handy, if a crust is to 
be pried up. The tapper's tools are few: a -Hnch steel bar, 3 feet 
long, a cast-iron ladle (6 inches in diameter, 3 inches deep), riveted 
to a 4-foot handle, a pick, a skimmer-rod to clean the well, a set 
of steel dies, a 4-pound hammer to number the bars with f-inch 
figures, and a chisel to trim them. The pot-pullers have to mop 
the pots with a clay wash ; they require no special tools, excepting 
occasionally a steel bar and sledge to loosen a cake of matte or slag 
from the slag-pot, should it adhere firmly. 

§ 65. Work on the Dump. — If the ordinary slag-pot is used, 
it is wheeled out on the dump and allowed to cool. When the 
contents have solidified the pot is tilted, and the cone of slag, with 
speise and matte adhering to the bottom, rolled out. This is broken 
up with a sledge, matte and speise are sorted out and piled up, and 
the clean slag is thrown over the dump. Sometimes the slag is 
loaded into flat cars in pieces as the result of breaking up the cones, 
or else the clean slag is first put through a crusher, thus preparing 
it for the direct use of the railroad as ballast. 

When the overflow-pot is used which discharges slag free from 
shots of matte into an ordinary slag-pot, it is allowed to cool on the 
dump. The contents of the other pot are emptied while still liquid 



SMELTING IN THE BLAST-FURNACE. 235 

over the edge of the dump. The scale of slag forming on its sides 
is reserved, if required for the blast-furnace. With a catch-pot 
having- a tap-hole, the pot is wheeled either to the edge of the dump 
and the liquid slag is discharged there, or to the large side-dumping 
pot, if this be in use. The tap-hole of the catch-pot is opened from the 
inside by thrusting a steel bar through the crust of the slag. The 
large tilting-pot is generally placed in a trench left open in the dump, 
and receives there from both sides the liquid slag. When filled, it 
is hauled to the edge of the dump by a horse or mule and emptied. 
Occasionally this receiving-pot is placed between two furnaces to 
receive their clean slag, and hauled, when filled, by a wire rope to 
the edge of the dump and there tilted. 

All the slag remaining in the overflow-pot or in the catch-pot 
with tap-hole goes back to the blast-furnace. 

The dragging-out of slag-pots to the edge of the dump is great- 
ly facilitated by keeping its surface even and smooth. This is 
done by forming squares (about three feet) with rails or sand on 
top of the rough surface and filling them with liquid slag. 

The number of men required on the dump varies with the size 
of the furnace and the manner of disposing of the slag. Two men 
in a ten-hour shift can clean up the slag and matte made by a 
furnace 42 by 120 inches at the tuyeres. This represents good 
work. The tools required are a steel bar, a pick, a sledge, a round- 
pointed long-handled shovel, and an iron wheelbarrow. 

§ 66. Irregularities in the Blast-Furnace.— The disturb- 
ances that occur during the run of a blast-furnace, having different 
sources, are numerous. Some are caused by defective machinery 
and apparatus, others by refractory ores, faulty charges, or wrong 
manipulations. Some accidents due principally to the last-men- 
tioned cause may be corrected in the following way. 

When the charges do not descend as evenly as they should, 
•one side sinking faster than the other, the jackets on the lower 
•side being much hotter and the tuyeres brighter than those on the 
upper side, the first change is made in the manner of feeding. The 
fuel is placed more on the hanging side, and ore and fluxes on the 
quickly-descending one ; then a full stream of water is turned into 
the hot jackets and that of the cool jackets reduced till there is 
just an overflow. By the combination of these remedies the smelt, 
ingof the furnace on one side more than the other will be corrected 
and the charges will right themselves again after a few hours. 
Shaking up the charge with a long, heavy (l£-inch) steel bar, 



236 METALLURGY OF LEAD. 

introduced through the feed-door into the hanging side, will often 
hasten matters. At some works that have to treat ores rich in 
zinc, it is the practice at the beginning of every day-shift to drive 
in a steel bar at the four corners of the furnace as far as the top of 
the jackets, thus loosening any wall accretions that are forming. 
By doing this, the number of times a furnace has to be barred down 
to remove accretions is greatly reduced and the whole running 
improved. 

When the charges descend irregularly, it often happens that the 
fire creeps up and it becomes hot on the surface (over-fire, fire-tops, 
hot top). The furnace may then be fed down, which consists in 
adding only just enough fresh charge to keep the flame or heavy 
smoke from passing through. When the surface has thus been 
lowered for 2 or 3 feet, the furnace is filled up quickly again and 
the top thus cooled. Simple sprinkling of water on the top of the 
charge has only a temporary effect. This feeding down helps mat- 
ters if the over-fire does not come from a crust in the crucible. 

The cause of the irregular descent of the charges lies generally 
in the formation of wall accretions (§ 82), which begin on top of 
the jackets and grow thicker towards the feed-door. They as- 
sume different forms. The following figures, 146-149, by lies 1 show 
some extremes. Figure 146 represents a more or less regular shape, 
and the smelting power of the furnace is not necessarily reduced. 
With irregular hangings, like those in Figures 147, 148, 149, the 
descent of the charges will be greatly obstructed and the amount 
of fluedust much increased. The charge will be tight at the nar- 
row parts of the furnace, and the blast entering the tuyeres will be 
concentrated in a few places and cause "blow-holes." As soon as 
these wall accretions are discovered they have to be cut or barred 
down. In order to reach the lowest part, the furnace is fed down 
and the blast lessened at the same time, till the charge has reached 
about the top of the jackets. While it is being lowered no lead is 
removed from the well, in order that the crucible may be entirely 
full while the barring out is going on. The blast is stopped, the 
blast-pipes are removed, the tuyeres closed, all the slag in the 
furnace is tapped, and the flow of water into the jackets nearly 
shut off. When everything is ready a charge of coke is given, wi*h 
some slag and flux, forming a bed for the accretions to fall on as- 
they are chipped from the walls. The cutting-out is best begun 

1 Engineering and Mining Journal, February 6, 1886. 



SMELTING IN THE BLAST-FURNACE. 



23' 



just above the jackets. A square-pointed steel bar about 1 J inches 
in diameter and long enough (about 16 feet) to reach from the top 
of the jacket well into the opposite side on the feed-floor, is driven 
with a sledge into the crust above the jacket. If it does not yield, 
a rope tied around the head of the bar is thrown to the opposite 
side, where several men pull it and thus break off the crust. This 
is repeated until it has been removed in a number of layers, two 
sets of men working on opposite sides. While the crust is being 
barred off, fuel, slag, and fluxes are added that it may be smelted 
out when the furnace is started up again. The reason that the 
barring down is begun from below is that otherwise the broken 
crusts and small slag-charges that have been added might so fill 





Fig. 146. 



Fig. 147. 





Fig. 149. 



WALL-ACCRETIONS. 

up the furnace that it would be impossible to reach the crust at 
the top of the jackets, and new accretions would form again quick- 
ly. Sometimes, however, if the crust is thick and hard, the bar- 
ring-down is begun from the top, continuing until the accumulated 
crust and slag-charge meet the clean side-wall, when it is smelted 
out, keeping the charges low. The furnace is now let down again, 
and the second half of the barring-down is begun at the top of the 
jackets and continued upward. 

When the sides of the furnace have been cleared, the tuyeres are 
cleaned out, the tuyere-pipes inserted, a weak blast, to be increased 
gradually, is turned on, and smelting resumed. Care must be taken 
about the water supply as the jackets grow hotter, and about the 
tapping of the slag, as the wall-accretions often melt very fast, and 



238 METALLURGY OF LEAD. 

there is danger of the slag's entering the tuyeres. After starting 
the furnace it is sometimes found that a small crust has formed 
over the lead while the blast was shut off. This is perforated 
with a long iron rod or with a steel bar, if necessary, and will soon 
disappear, if the furnace was in good working order below before 
the cutting-out began. 

A second method of barring down a furnace, said to work well, 
is to let down the charge to below the upper rim of the jackets, 
and give a bed of coke as previously described. In the meantime, 
the bricks between the jackets and the collar, on which the shaft 
rests, are removed for a distance of 1 or 1^ feet. The accretions 
are cut down, beginning from above, and raked out through this 
opening. When the shaft is clean, the opening is bricked up, light 
charges are given, the furnace is filled up with ore-charge, and the 
blast let on. The accretions go to the roasting furnaces. 

A third method of barring down, applied in very bad cases, is 
the following ; it is dangerous, but may be necessary if the life of 
the furnace is in jeopardy. The charge is let down to the top of 
the jackets, and cutting out begun from the top and continued un- 
til the crust collecting in the furnace has risen so high that the 
work cannot proceed any further. The two breast-jackets are now 
taken down and the contents of the furnace raked out (§ 67), in 
order that the cutting may continue until the walls are clean. This 
may take as much as eighteen hours. Two parties of three men 
each, working on opposite sides of the furnace, do the cutting, one 
man holding the bar and two sledging. As the work is hard and 
has to be done as quickly as possible, the regular hands have the 
constant assistance of furnace-men and helpers from other furnaces, 
who work half an hour at a time. While the accretions are being 
removed, a crust 6 inches thick or less will form on top of the lead. 
This is broken up, the breast-jackets are put back, and the lower 
front is closed. Charcoal is fed from above to the top of the 
jackets and the furnace blown in anew, adding twenty-five bars of 
lead to the first slag charges to heat up the lead in the crucible. 

Guyard * suggests that by using caustic lime in the charge in- 
stead of limestone, these accretions might be made less troublesome, 
as lime has a decomposing action on the sulphides, of which they 
consist in great part. The writer is not aware that caustic lime is 

1 Emmons, "Geology and Mining Industry of Leadville," pp. 728 

and 747. 



SMELTING IN THE BLAST-FURNACE. 239 

used in any lead furnace. It would seem as if the lime, being ex- 
posed during its descent in the furnace to gases rich in carbonic 
acid, would be converted quickly into carbonate of lime in the 
upper part of the furnace and lose any decomposing power. Again, 
while accretions next to the furnace walls consist mostly of sulphides 
they are usually covered by a thin crust as hard as flint ; this is 
followed by a softer substance that is often a powder, and this 
again covered by a crust so tough that it is sometimes extremely 
difficult for a steel bar driven hard with a sledge to produce any 
effect on it. It does not seem probable, therefore, that caustic 
lime would have any important effect on these composite crusts, 
which must differ from sulphides as much in their chemical prop- 
erties as they do in their physical. 

It often happens, even when a furnace is otherwise doing well, 
that the tap-hole becomes hard and the tuyeres dark. The fuel is then 
distributed more over the cold places, additional fuel being given 
for a short time if necessary. One or two bars of lead are some- 
times placed on top of the charge over the tap or the tuyere, but 
this cannot be commended, although it is often effectual. The 
change in feeding the fuel will generally soften the tap. To soften 
a crust in front of a tuyere, the bag or the gate should be closed, or 
nearly closed, as the blast playing on chilled slag can only have a 
bad effect. If it is turned off, the heat generated at the neighbor- 
ing tuyeres will melt off the crust ; then a little blast is allowed to 
pass through the tuyere, and gradually increased until the full 
blast can be turned on again. 

This proceeding may also be necessary if some fine ore trickles 
through the coarser parts of the charge, appearing in a crude state 
at the mouth of the tuyere, and, as is sometimes the case, runs into 
the tuyere pipe. If fine ore appears at the same time at several 
tuyeres and too many would have to be tied up, the easiest remedy 
is to feed down the furnace and thus loosen the charge. If this 
docs no good, a coarse ore-charge will have to be substituted for a 
short time, or if coarse ore is not to be had, a slag-charge. 

The forming of hearth accretions (§ 83) is indicated on the 
feed-floor by the top becoming hot and the charges not sinking 
regularly, but in jerks, a foot at a time, and blazing up with every 
settling. On the furnace-floor the lead in the well becomes dark, 
and does not play with the blast. By inserting a rod through the 
tap-hole, tin; position and often the thickness of the crust can be 
ascertained. Sometimes it only forms a ridge across the furnace, 



240 METALLURGY OF LEAD. 

and communication is open in front and at the back; more com- 
monly the crust begins at the back and grows toward the front, 
gradually closing all communication. The lead produced in the 
furnace cannot get into the crucible, and runs out with the slag. 
A case like this generally needs a change in the composition of the 
charge; the slag analysis will show the defect and give the remedy. 
As it takes from six to eight hours before a new slag makes its ap- 
pearance, holes are driven with a steel bar through the crust that 
the lead may find its way into the crucible. If the crust has not yet 
hardened, it is repeatedly perforated by introducing an iron rod 
and working this up and down and to both sides. The crust may 
be only a temporary affair, and can then be worked into the slag, 
which removes it. Cases do occur where the crust will not yield 
quickly enough to being fused out and has to be removed by force. 
This has been ironically spoken of as "muscular smelting," but it 
is sometimes unavoidable. Before beginning, all the slag in the 
furnace is tapped, the basin of the lead-well covered with glowing 
charcoal, the blast stopped, the tuyere-pipes drawn out, the tuyeres 
closed, and the flow of water into the jackets reduced. The 
breast of the furnace is now removed wholly or in part, any 
chilled slag in front is chipped off, and some of the loose material 
raked out into a wheelbarrow in front of the furnace and taken 
away. A heavy steel bar is passed through two opposite tuyeres 
nearest the breast to hold up the charge. Balls of loam or clay 
are tamped behind the lower part of the front jacket to prevent 
the charge from rolling down. The balls, placed on the end of a 
board, are slid below the front of the jacket and rammed upwards 
against the charge with a rod bent to a hook. When this is done, 
any lead that has accumulated in front is ladled out, and the crust 
thus laid bare. A hole is driven through it with a heavy steel 
bar. It is necessary to have a number of these read}', as the 
points soon become dull or bent. The hole is enlarged by driving 
the bar again close to it and breaking the crust towards it. When 
large enough to receive a 2-inch steel bar, this is warmed, inserted 
into the hole, the crust pried up, and the broken pieces raked out. 
If the crust will not yield, the hole is enlarged and the furnace 
started up again. The lead previously bailed out is returned, and, 
if necessary, fresh hot lead added, the clay balls are removed, the 
hollow space in front is filled with charcoal, the steel bars are with- 
drawn from the tuyeres, the breast is put in, the tuyeres are 
opened, the pipes inserted, and the blast is let on, but very gently 



SMELTING IN THE BLAST-FURNACE. 241 

at first. When the first two or three pots of slag have been tapped 
the rod is repeatedly inserted to keep the hole open until the new 
slag comes down. The lead will soon show the effect of having 
communication between it and the slag partly restored; it be- 
gins to play with the blast and becomes hotter, thus assisting the 
work of the new charge. 

A crust is sometimes caused by a leaking jacket. This is first 
indicated by the appearance of moisture at the tuyere or the bot- 
tom of the jacket. The leak, if small, can be temporarily stopped by 
mixing corn-meal with hot water, pressing it with the hand into 
small balls, and throwing these into the water-feeder of the jacket. 
Soon, however, the jacket will have to be removed. For this pur- 
pose the crust, on the inside is first allowed to grow thick by cool- 
ing, which is done by turning in a full stream of water, and open- 
ing the discharge at the bottom. Two courses of brick are chiselled 
out above the jacket. When cool, the furnace is stopped, the 
cooling-water on the side of the injured jacket shut off, the water- 
trough removed, and the injured jacket unhinged, taken out, and a 
new one put in its place. The whole procedure need not take more 
than twenty minutes. Should the crust on the inside of the jacket 
prove too thin and breakout, the opening is closed by the introduc- 
tion of clay balls. The space where the new jacket is to be insert- 
ed must be absolutely clean, as any little pieces of brick or other 
hard matter will obstruct the placing of the new jacket and cause 
much delay. Of course the foregoing has no reference to a fur- 
nace with wrought-iron jackets extending its entire length. 

The clogging up of the lead-well has still to be considered. In 
smelting sulphide ores rich in lead, sulphide of lead held in solu- 
tion in the crucible often separates out when the lead ascends the 
channel toward the basin. A bent iron rod may be inserted to 
clean it out. This presence of sulphides is generally caused by an 
incomplete decomposition of galena in the furnace on account of 
lack of heat in the smelling zone. If the charge is rich in copper, 
this causes coppery lead gradually to close up the channel. 

§ 07. BlOWillg-Ollt.— If the furnace needs to be repaired, or 
if an accident happens that cannot be remedied in a short time, 
say eighteen or even twenty-four hours, the furnace has to be 
blown out. This is done by stopping the ore-charges and substi- 
tuting slag charges, until most of the ore lias been smelted out. 
The charge is allowed to sink and the blast is gradually lowered. 
Soon volumes of dark smoke mixed with white lead-fumes will 



242 METALLURGY OF LEAD. 

appear. "When the charge has receded somewhat, and before a 
iiame makes its appearance, the damper in the flue leading to the 
dust-chamber is closed, and the fumes are conducted into the air by- 
opening the damper on the top of the furnace or by lowering the 
sheet-iron stack, or by whatever contrivance may be in use for the 
purpose. If this were not done, an explosion might occur in the 
dust-chamber. To check the flame and to reduce the temperature, 
water is often sj:>rinkled over the charge, although its effect on the 
lining of the furnace cannot but be deleterious. AVhen the charge 
has sunk as far as the top of the jackets, the blast is stopped and 
the tuyere-pipes are removed. All the liquid slag is tapped, the tap- 
ping-jacket removed, and the breast of the furnace is knocked in. 

Sometimes the furnace is blown down, allowing the charge to 
sink only till heavy fumes, but no flames, appear and the entire 
contents then drawn. In this case there is no need of closing the 
damper to the dust-chamber ; in fact, many furnaces have no 
damper at all. 

The bulk of the slag remaining in the furnace is withdrawn 
with a hoe into iron wheelbarrows, which are emptied on the dump 
and then chilled with water. As it is important that there should 
be little delay in drawing the charge, four or five wheelbarrows are 
placed one behind the other near the front of the furnace. As 
soon as the first is filled with red-hot charge it is wheeled away 
and replaced by the second, the emptied wheelbarrow being put at 
the end of the line. When all the charge that can be easily reached 
with the hoe has been drawn out, the front jackets are taken down 
and the rest removed. Meanwhile a thin crust will have formed on 
top of the lead in the crucible. This is easily broken, and the lead is 
then ladled out into the moulds that have been moved from the 
lead-well to the front of the furnace. 

§ 68. Furnace Books. — A daily record of the work done is 
kept for every furnace. The following is one of the many suitable 
skeletons (see Table A, page 243). 

The time when any change in the charge is made, or when any- 
thing out of the regular way occurs is noted under the first column 
of " Remarks." The second column of " Remarks " refers to the 
shipment of the bullion. 

§ 69. Furnace-Assay Book. — A separate book is kept to re- 
cord the assays and analyses made of slag and matte (see Table B, 
page 243). Under the head of "Remarks" are brought the names 
of any of the other furnace-products that may be assayed now and 
then. 









0& 



u O 

- 



.a o 



OcL 



-- g 



&» 



C 5' 



£ * s 






■ a 
E« 



244 



METALLURGY OF LEAD. 



FURNACE PRODUCTS. 

§ 70. Furnace Products. — The products of a blast-furnace 
are base bullion, speise, matte, slag, wall accretions, hearth ac- 
cretions, furnace cleanings, and flue-dust. 

§ 81. Base Bullion, the commercial name for argentiferous 
lead, as distinguished from silver or gold bullion, contains in ad- 
dition to the silver and gold small quantities of other metals, as seen 
by the following analyses: 





Clausthal. 


Sophien- 
htttte. 


Freiberg. 


Mecher- 
nich. 


Pribram. 


Leadville. 


Leadville. 


Pb 


98.80944 


99.641096 


95.088 


99.5913 


97.3597 


99.0798210 


98.492379 


Ag 


0.1412 


0.000250 


0.470 


0.0215 


0.4230 


0.6112445 


0.793417 


Bi 


0.0048 


0.352053 


0.019 




0.0070 


trace 


0.011791 


Cu 


0.1862 


0.000279 


0.225 


6.1332 


0.1100 


0.0479100 


0.071450 


< Od 


trace 




none 


.... 




trace 


trace 


As 


0.0664 




1.826 




0.2900 


? 


? 


Sb 


0.7203 


0.002872 


0.958 


6.2180 


1.5240 


0.2138940 


0.347881 


Sn 




.... 


1.354 




0.2500 


trace 


0.000897 


Au 






.... 






0.000888 


0.000891 


Fe 


0.0664 


0.002877 


0.007 


0.0300 


0.0036 


0.0063000 


0.012600 


Zq 


0.0028 


0.000573 


0.002 


0.0060 


0.0012 


0.0016052 


0.000232 


Ni 


0.0023 


.... 


.... 


trace 


0.0015 






i Co 


0.00016 






.... 


.... 






! S 







0.051 




0.0300 




0.048934 


Reference : 


l. 


2. 


3. 


4. 


5. 


6. 


6. 



1, Zeitschrift fiir Berg-, Hiitten- u. Salinen-Wesen in Preussen, xviii., p. 303; 2, Zeitschrift 
fur Berg-, Hiitten- u. Salinen-Wesen in Preussen, xix., p. 169 ; 3, Wagner's Jahresberichte, 1887, p. 
401 ; 4, Berg- u. Hiittenm. Zeitung, 1886, p. 434 ; 5, Oesterreichisches Jahrbuch, xxxix., p. 52 ; 6, 
Emmons, " Geology and Mining Industry of Leadville," p. 694. 

The impurities of a bar of lead, with the exception of silver 
and gold, always collect nearer the top than the bottom. This is 
illustrated by the analyses of Streng, made from non-argentiferous 
lead, and bv Schertel, made from base bullion. 





Streng.* 


Schertel. + 


Top. 


Bottom. 


Top. 


Bottom. 


Top. 


Bottom. 


Ag 

Bi 


3. '621 
6.274 

0.008 

0.003 
0.148 


1.242 
6 .158 

o'.'oos 

trace 

0.082 


6.508 

0.090 

6*012 
0.002 
0.012 


6.'l40 

6.057 

6 008 
trace 
trace 


0.423 
0.132 
1.324 
2.164 
0.700 
0.941 
0.103 
0.016 
0.029 
0.500 
10.321 


0.403 
0.042 
0.034 
1.980 
0.749 

? 
0.009 
0.003 

10.824 


Cu 


As 


Sb 


Sn 


Fe 


Zn 


Ni 


S 


Sn. err 





* Berg- u. Hiittenm. Zeitung, 1859, p. 14. + Wagner's Jahresberichte, 1887, p. 401. 



SMELTING IN THE BLAST-FURNACE. 245 

Schertel's analyses would seem to show that silver has a similar 
tendency. He, however, took his sample from lead that had been 
kept for twenty-four hours above its melting-point in an iron 
cylinder, 3 feet 3| inches high, and thus created special conditions. 
In bars of base bullion which cool quickly just the reverse is the 
case. The following tests made by the writer in 1881 may sub- 
stantiate this. (Figures 150 and 151.) 

Bars of base bullion were sawed into three pieces and samples 
taken as shown by the numbers and assayed separately. 

(1.) 149.7 (2.) 150.5 (3.) 146.0 (4.) 148.7 (5.) 145.0 

Fig. 150. ■[ (6.) 137.0 (7.) 152.0 (8.) 149.0 (9.) 149.0 (10.) 151.0 

(11.) 148.0 (12.) 150.0 (13.) 150.5 (14.) 150.0 (15.) 152.0 



(1.) 127.0 (2 ) 134.5 (3.) 128.5 (4.) 129.5 (5.) 125.0 

Fig. 151. \ (6.) 133.5 (7.) 124.0 (8.) 132.0 (9.) 126.5 (10.) 134.5 

.) 129.0 (12.) 132.0 (13.) 133.0 (14.) 132.0 (15.) 134.0 



( Cl.] 

A (6.; 

((HO 

The results (ounces i>er ton) clearly prove that the whole lower 



FIG.150SB ? 8 D^ a b c 2 F 4 Vh 101 FiG.151 

SECTION* 5 11 Ofhi " ~2 8ECT10 NaH. 

A B. SECTION e SECTION 

c o. FlG.152 

BARS OF BASE BULLION. 

part of the bar is richer than the upper and that the centre of the 
upper surface is the poorest part. Similar results were obtained by 
Piquet. 1 

This concentration of the silver in the lower part of the bar can 
be explained 2 by the separation of the argentiferous lead, while 
cooling into crystals low in silver and liquid lead high in silver. 
When a bar has been moulded, the surface cools first, and crys- 
tals begin to form there; then the sides slowly solidify, and a mass 
of Liquid Lead and crystals remains in the centres. As the cooling 
proceeds from the to]) downward the crystals will continue to form 
on the cooler upper side and gradually force the richer liquid lead 
towards the bottom. The reason that the surface of a bar is so 
much poorer than the other parts is not only because of the crys- 
tallization, but also on account of the impurities that rise to the 
surface when a bar is cooling, forming the dross. They run 
much lower in silver than the pure lead. If bullion is so rich in 

1 Roswag, " DIsargeutatioD du plomb," Paris, 1884, p. 127. 

2 Roswag, Op. cit., p. 120. 



246 



METALLURGY OF LEAD. 



dross that the lead cannot eliquate freely, the natural distribution 
of the silver will be much disturbed. This will account for the 
fact that assays from the top of a bar run sometimes higher than 
those of the bottom. For instance, coppery bullion from the 
Ramshorn silver mine, Idaho, containing so much dross that, if 
dropped on the floor, it would break, gave to Rhodes 1 the follow- 
ing results in ounces per ton (Figure 152): 



a. 432 
d. 450 
g. 385 



b. 432 
e. 434 
h. 397 



c. 439 
/. 441 
i. 386 



According to Kempf, Nenninger, & Co., 2 gold also seems to be 
concentrated with the silver near the bottom of the bar. 



Sample from 


Ag. 


Au. 


Ag. 


Au. 


Top 

Bottom 


143.3 

148.4 


4.59 
4.76 


129.7 
134 .3 


3.68 

3.82 



Similar results have been published by Torrey and Eaton. 3 
§ 72. Sampling and Assaying. — The irregular manner of 
sampling and the unequal distribution of the silver have caused 
much trouble in estimating the value of base bullion. Samples 
used to be taken, for instance, with a gouge from top and bottom 
near the opposite ends of a bar. This method has two sources of 
error, the form of the sample and the fact that they are taken only 
at the ends. The sample obtained is a conical chip having its base 
at the surface and its apex % inch or less below. The incorrect 
form has been remedied by using a punch which gives a cylindrical 
chip about -§■ inch in diameter. If driven half way through the bar, 
say 2 inches, it will represent a correct sample of that part of the 
bar. Sampling, however, from the ends of the bar gives too high 
a result, as can be seen by turning to Figures 150 and 151. The 
poor parts of the bar, represented by No. 6 in Figure 150 and 
No. 5 in Figure 151, are excluded. The only correct method which 
is simple and quick is to take punch samples diagonally across a 
row of, say, five bars by driving the punch every time deeper than 
half way through, turning over the bars and repeating this on the 
other side in the opposite diagonal. Three or four blows with a 

1 Private communication, July, 1891. 

2 Engineering and Mining Journal , July 1, 1882. 

3 Ibid., December 25, 1886. 



SMELTING IN THE BLAST-FURNACE. 247 

four- or five-pound hammer are sufficient to drive a stout punch 
having an -^-inch opening about 2 inches deep; a few light taps on 
the sides loosen the punch and break the chip. 

Austin l recommends removing from a bar 4 inches thick a 
chip a little longer than 2 inches, slipping it into a hole bored 2 
inches deep into a block, and trimming off the projecting end 
with shears. This is too troublesome and is not followed. While 
in the method previously described a few chips may be a little 
longer than 2 inches, others are a little shorter, and the average chip 
will have the desired length. 

Drilling one or more holes through every bar and taking the bor- 
ings as sample has been suggested. The writer is not aware that 
this method is carried out anywhere. The large bulk of the sam- 
ple and the liability to loss have probably prevented its use. 

At some works a drill has lately been introduced which removes 
a cylindrical chip of the desired length from the bar. The opera- 
tion requires no hard work and is quick ; the bit has no compli- 
cated construction and outlasts many punches ; the whole appara- 
tus is simple and light, so as to be easily moved from one part of 
the works to the other. 

The chips obtained by* sampling a car-load (weighing from ten 
to fourteen tons) or a lot are collected in a wooden box, melted 
down, and cast into a bar. The writer 2 has always obtained sat- 
isfactory results by the following procedure. The chips are melted 
down quickly in a graphite crucible, which has first been heated for 
some time, so as to be dark-red near the bottom. When melted 
and heated till the dross is about redissolved, but before any cupel- 
ling has begun, the lead is well stirred with an iron rod for several 
minutes and then poured with the dross into a mould of, say, 3 by 
7 inches. This sample-bar, representing from ten to fourteen tons 
of base bullion, weighs about eight pounds, rather more than less. 
Another form of mould is 4 by 12 inches, which gives a thinner 
bar. It has a ridge on the longer centre-line. The depression this 
makes m the bar marks the place where it is cut, if it is necessary 
to divide it in order that hali may be sent to the buyer. 

A different method of melting the chips is followed by Kerapf, 
Nenninger & Co. 3 They call attention to the fact that the chips 

from the top of the bar are liable to be smaller than these from 

1 Engineering and Mining Journal September'.), isso. 

" " Mineral Resources of the United states," L888 84, p. 464. 

8 Engineering and Mining Journal, July 1, 1882. 



248 METALLURGY OF LEAD. 

the bottom, on account of the hardness of the dross and the un- 
evenness of the surface. They melt down top and bottom chips 
separately, take from the resulting two bars an equal amount, 
and melt into a second bar, their final sample-bar. 

At some smelting-works they hold back the poor dross (in 
pouring) with a rod, casting only the liquid lead, and use this as 
the sample. Refining-works, of course, object to this. If the dross 
is held back, as "Refiner" 1 suggests, and practises at his works, 
it must be assayed separately from the liquid lead. The final as- 
say is calculated with reference to the weights of dross and lead 
and the result obtained is good : 

A.a + B.b . 
— -r ^ — = Average assay. 

A. = weight of dross ; a = assay of dross. 

B. = " " lead; b= " " lead. 

While at the works of "Refiner" the writer practised this 
method. Where a large quantity of base bullion is desilverized, 
it takes too much time. With bullion rich in dross "Refiner's" 
method is to be greatly recommended. 

From the 3 by 7-inch sample-bar four samples are taken from 
the middle of the sides, cutting the entire bar through. From the 
4 by 12-inch bar the samples are taken from half of the bar only : 
two on the outer side and two along the central depression. The 
w T riter prefers to cut out samples weighing a little over one assay- 
ton and to use four samples of one assay-ton each for the sub- 
sequent cupellation, as his results have been more accurate and 
uniform with silver than when half an assay-ton or less lead 
has been taken. That the results with gold would be more satis- 
factory from four one-assay ton assays than with a smaller amount, 
especially if there is little gold, as is generally the case, is clear 
from the advantage of the added weight, when the silver buttons 
have been dissolved in nitric acid. 

The assay sample of bullion is usually cupelled without being 
scorified at as low a temperature as possible. Every cupel must 
show feather litharge, if the assay is to be accepted as definite. 
To regulate the temperature automatically Torrey 2 has constructed 
a heat-regulator, but generally the assayer uses his judgment as to 
degree of heat and draught of his muffle. If the temperature is 

1 Engineering and Mining Journal, May 20, 1882. 

2 Engineering and Mining Journal, August 28, 1886. 



SMELTING IN THE BLAST-FURNACE. 



249 



too low toward the end, lead is liable to remain with the silver, 
often also copper and bismuth. With impure bullion it is advisable 
to test the nitric acid solution, obtained in dissolving the silver 
buttons when determining the gold, for lead with sulphuric acid 
and for copper and bismuth with ammonia. Impure base bullion 
has to be scorified before it is cupelled, especially if it contains 
much dross (arsenic, antimony, and copper). Part of a table of the 
losses endured in cupelling argentiferous lead lately published by 
Rossler 1 is subjoined. 



Material. 


Ounces Ag 
per ton. 


Charge. 


Multiple of 
Pb. 


Loss of Ag in 
percentage of 
amount pres- 
ent. 


Mgrs. Ag used. 


Grs. Pb. 


Argentiferous 

copper 

Argentiferous 

copper 

Argentiferous 

lead 

Argentiferous 

lead 

Argentiferous 

lead 

Argentiferous 

lead 

Argentiferous 

lead 

Argentiferous 

lead 


43.75 

583.32 

43.75 

87.49 

218.74 

437.49 

874.98 

4374.90 


10 mgrs. Ag + 
10 grs. Cu. . . 

200 mgrs. Ag+ 
10 grs. Cu... 

150 

<« 

tt 

n 
tt 

tt 


200 

160 

100 

50 

20 

10 

5 

1 


12,000 

800 

600 

300 

120 

60 

30 

6 


8.3 
4.5 
2.5 
2.2 
2.0 
1.6 
0.9 
0.4 



£ 73. Speise. — The speise obtained in lead-smelting is princi- 
pally an arsenical speise. The predominant element is iron ; this is 
somewhat replaced by nickel, cobalt, copper, and to a small extent 
by lead, bismuth, gold, and silver. 

Guyard calls attention to the absence of cobalt in Leadville 
speise, which he found concentrated in the dross skimmed from the 
lead-well. He also found as much as 10 per cent, of grains of 
metallic iron free from arsenic in Leadville speise, which is uncom- 
mon. It has been already stated that speise always contains shots 
of lead, and thai when coarsely-crystalline less is found than when 
it is fine-grained. As regards the presence of precious metals, the 
fact is to be noted that speise retains considerable amounts of 

1 Berg-u. Huttenm. Zeitung, 1888, p. 480; Engineering and Mining 
Journal, July 36, L890. 



25U 



METALLURGY OF LEAD. 



gold, while very little gold is found in matte ; speise-assays show 
from a trace to 0.5 ounces gold per ton. 

To treat speise so as to extract the silver, gold, and copper 
economically has always been a difficult problem. In a large 
number of works it is not treated at all, as only small quantities 
are produced. The cheapest and simplest way is to roast it in a 
heap of about fifty tons, which burns from two to four weeks. 
The imperfectly-roasted speise is sorted out, crushed and roasted 
in a calcining furnace. The whole is then smelted in the blast- 
furnace with pyrite or matte. The result will be base bullion and a 
matte rich in copper and silver, and perhaps a small amount of speise, 



As... 
Au . . . . 
Cu . . . . 

Pb 

Mo 

Fe 

Zn 

Ni 

Co 

S 

As . .. 

Sb 

SiOo . . . 
CaO . . . 

Reference 



' Leadville. 


Leadville. 


0.0085 


0.0301 


trace 


0.0009 


0.3628 


0.2566 


1.4935 


2.5030 


0.2110 


0.2155 


60.5780 


70.4780 


trace 


trace 


0.0876 


0.0981 


5.8191 


4.4695 


31.4725 


21.8003 


trace 


0.1450 


l. 


l. 



Pueblo. 



Pueblo. 



Eureka j Pribram. Pribram. 



up to 0.016 up to 0.014 ) Q29 



2.09 

1.87 

58.32 

trace 
trace 

4.105 
30.005 
trace 

0.15 



5.06 
0.69 

59.42 

trace 
trace 
2.80 
31.17 
trace 
trace 



1.06 
2.18 
2.31 
57.02 
0.07 



3.34 
32.95 
0.13 
0.23 
0.34 



3. 



0.037 

1\956 
1.752 

61 '.330 

2! 056 
0.194 
9.600 
18.750 
2.450 

Ca 6.500 

4. 



0.020 

0. 409 
3.245 

56 '.700 

I 0.783 
'10. 000 

26.757 
1.608 



Ca 0.535 



1, Emmons, "Geology and Mining Industry of Leadville,"' p. 720; 2, Dewey, Bulletin No. 42. 
''United States National Museum," p. 52 ; 3, Curtis, "Silver-Lead Deposits of Eureka, >~ev..'" Mono- 
graph vii., United States Geological Survey, 1884, p. 160 ; 4, Balling, Berg- u. Hiittenm. Ztg., 1867, 
p. 419. 

in which any nickel and cobalt will be concentrated. This second 
speise goes to a new heap of first speise, as nickel and cobalt occur 
in such small quantities as not to call for any further attention. 

Davies 1 invented a process for desilverizing speise which is said 
to give satisfaction at the Eureka Consolidated Works, Eureka, 
New It consists in tapping 800 j)ounds speise into a small cyl- 
indrical iron converter lined with fire-brick (Figure 153), adding 
from 20 to 25 per cent, liquid lead and introducing from the 
bottom a blast of 17 ounces pressure through a 4;-inch pipe 
for 3 or 4 minutes. This stirs up the lead and speise and 



Engineering and Mining Journal, June 30, 1888. 



SMELTING IN THE BLAST-FURNACE. 



25L 



burns off some arsenic. Most of the silver and gold is taken up by 
the lead ; the liberated iron corrodes somewhat the lining, but eats 
through only very slowly. The converter is turned down and the 
contents are discharged into a cast-iron receiver having the form 
of a slag-pot, the bottom of which has a £-inch hole for tapping 
the lead. The desilverized speise solidifies quickly, and when hard 
the still liquid lead is tapped. Best results are obtained when the 
lead assays 40 ounces to the ton. The desilverized speise is then a 




Fig. 153.— converter for desilverizing speise. 



waste product. The claim is made that from a speise of the compo- 
sition shown in the table, 85.5 per cent, of the silver and 89.28 per 
cent, of the gold are extracted, with a loss of lead varying from 5 
to 8 per cent. 

§ 74. Analytical Determinations.— The metals to be de- 
termined by assay in raw or roasted speise are silver and gold 

(§ 34), lead (§ 79), copper, arsenic, and iron (§ 8 1). 

l. Copper.— In determining the copper by titrating with po- 



252 METALLURGY OF LEAD. 

tassium cyanide, the presence of arsenic ' has an injurious influence. 
To counteract this, Blake 2 recommends two methods, both of which 
give satisfactory results. One of them only is given here. The 
weighed sample is heated with sulphuric acid (concentrated acid 
diluted with an equal volume of water); then nitric acid is added 
to complete the decomposition. The solution is evaporated and 
heated half an hour after the sulphuric acid fumes have appeared; 
it is then diluted, boiled, and the copper titrated after supersatu- 
rating with ammonia, either directly or after having been first 
precipitated with zinc and redissolved. (For details of the cyanide 
assay see § 69.) 

2. Arsenic. — The quickest assay-method for arsenic is the one 
by Pearce. 3 Half a gramme or less (according to the amount of 
arsenic present, as 1 part (Ag 3 As0 4 ) 2 weighs nearly six times as 
much as 1 part As) of the finely-pulverized substance is melted in 
a large porcelain crucible with from 6 to 10 times its weight of a 
mixture of even parts of sodium carbonate and potassium nitrate 
kept for five minutes in a state of fusion, cooled, and treated with 
warm water, which dissolves the alkali arseniate formed. The solu- 
tion is filtered,the filtrate, acidulated with nitric acid, boiled to drive 
off carbonic and nitrous acids, cooled, and carefully neutralized by 
first adding a slight excess of strong ammonia, then acidifying 
again by adding a few drops of concentrated nitric acid, and finally 
adding dilute ammonia drop by drop until it takes litmus-paper 
half a minute to show a change of color. 

In determining the arsenic of ores, some alumina may be pre- 
cipitated, which is filtered off. 

Canby 4 adds an excess of an emulsion of zinc oxide instead of 
neutralizing carefully with ammonia, and titrates the silver as 
shown below. 

The solution is stirred vigorously, the arsenic precipitated by 
adding an excess of a neutral solution of silver nitrate, as brick-red 
silver arseniate (Ag 3 As0 4 ) 2 , the liquid is filtered, and the precipi- 
tate washed with cold water. To determine the silver, two meth- 
ods may be employed : (a) The silver arseniate is scorified and the 
resulting lead button cupelled ; (b) the silver arseniate is dissolved 
in nitric acid and titrated with ammonium sulpho-cyanide, using 

1 Torrey and Eaton, Engineering and Mining Journal, June 6, 27, 188o. 

2 Trans. A. I. M. E., ix., p. 317. 

3 Engineering and Mining Journal, May 5, 1883. 

4 Trans. A. I. M. E., xvii., p. 77. 



SMELTING IN THE BLAST-FURNACE. 



253 



5 grs. NH 4 SCN to the liter and 5 cc. of a saturated solution of 
ammonium ferric-sulphate as indicator. 

1 part Ag = 0. 23148 parts As. 

The results are accurate. Antimony is without effect, while 
molybdic and phosphoric acids interfere with the method. 

§ 75. Matte. — The matte produced in lead smelting is princi- 
pally a comj)ound of iron and sulphur, in which the iron has been 
in part replaced by lead and copper and in a less degree by zinc, 
silver, nickel, cobalt, manganese, arsenic, antimony, calcium, ba- 
rium, and magnesium. 





Clausthal, 
old. 1 


Clausthal, 
new. 2 


Pribram. 3 


Pueblo. 4 


Pb 

Cu "... 

As 

Sb 

Sn 


41.50 
0.36 

0.68 

6.12 

34.05 

23.82 


11.399 
5.976 

0.212 

0.0311 
51.963 
1.963 
0.300 

0.112 

0.261 
0.706 

0.235 

25.807 


11.16 
1.59 
0.55 
0.93 

trace 
0.105 

41.31 

11.55 

1.40 

trace 

trace 

3.06 

0.05 

trace 

trace 

22.23 

4.79 


10.72 
0.61 
0.56 

none 

0.084 
52.27 
4.27 

0.41 
0.47 

24.015 


Ag 

Fe 

Zn 

Mn 


MnO 

Ni 

Co 

SiO., 


Ca .". 

CaO 

Mg 

Al 2 3 

S 


.COO 



1 Balling. " MetallhUttenkunde," 1 p. 86. 

2 Private notes. 

3 Oesterreichisches Jahrbuch, xxxix., p. 24. 

* Dewey, Bulletin No. 42, "United States National Museum, 1 ' p. 5£ 



The very high percentage of lead in the matte produced for- 
merly (1807) in the Upper llartz was connected with an enlarge- 
ment of the furnace at the zone of fusion, made to lower the tem- 
perature and thus reduce the volatilization of lead and the corrosion 
of furnace-walls. This resulted, however, in an incomplete decom- 
position of galena. In modern furnaces that are contracted at the 

Smelting-znne a matte contains from 8 to 12 per cent, of lead, if 

not otherwise enriched, e.g. y by the presence of zinc in the charge. 

It often happens that, if the constituents of a matte are figured as 



254 METALLURGY OF LEAD. 

sulphides, the analysis does not show enough sulphur to combine 
with the metals as FeS, PbS, Cu 2 S, NiS, etc. The explanation 
given by Rammelsberg 1 for lead sulphide, and by Minister 2 and 
Schweder 3 for iron and nickel sulphide, is that, as subsulphides do 
not exist, lead, iron, and nickel are held in solution by their sul- 
phides while liquid, and separate out during solidification. Mack- 
intosh* claims the existence of a subsulphide of nickel and iron. 
The value of a matte depends on the amount of silver, gold, lead, 
and copper, and (in exceptional cases) of nickel and cobalt it con- 
tains. Matte does not retain much gold, but considerable silver 
(§ 7). This is illustrated in a very interesting way by the fact 
observed by Carpenter, 5 that if a gold and silver bearing iron 
matte produced in pyritic smelting be added raw to the charge 
of a lead blast-furnace, all the gold, and from 60 to 70 per cent, of 
the silver, will be taken out. The first matte produced in a lead 
furnace rarely contains more than 5 per cent, copper and from 1 to 
3 per cent, nickel-cobalt. 

The object aimed at in working the matte is to extract the silver 
and gold by means of lead, and to concentrate the copper into an 
enriched intermediary product, which is sold to copper works. In 
exceptional cases it is sometimes worked at large refineries into 
metallic copper or blue vitriol. The two operations by which matte 
is concentrated consist of roasting and smelting, with an acid flux 
to slag the iron and a lead ore to take up the silver and gold. 

§ 76. Roasting of Matte. — Roasting is carried on in heaps, 
stalls, kilns, and reverberatory furnaces. The choice of method will 
depend greatly on the richness of the matte in silver. In roasted 
matte, according to Plattner, 6 the exterior part of each piece or 
pellet contains more silver and lead than the interior. Any dust 
that is made in handling the roasted matte, coming from the sur- 
face, will cause considerable loss in silver and lead ; again, if ex- 
posed to the rain, silver sulphate formed in roasting is liable to 
be lost by leaching ; finally, with rich matte the aim must be to 
roast as quickly as possible, so as not to lock up too much capital 
for months at a time. 

1 " Lehrbuch der chemischen Metallurgies Berlin, 1865, pp. 48 and 232. 

2 Berg- u. Huttenm. Zeitung, 1877, p. 195. 

3 Ibid., 1879, p. 18 ; Iron, xiii., p. 292. 

4 Trans. A. I. M. E., xvi., p. 117. 

5 Private Communication. May, 1892. 

6 " Die metallurgischen Rostprocesse," Freiberg, 1856, p. 205. 



SMELTING IN THE BLAST-FURNACE. 255 

1. Roasting in Heaps has the advantage of cheapness and 
simplicity of plant ; the matte need not be crushed fine, and most of 
the roasted matte is obtained in lump form, which is desirable for 
the blast-furnace. This is, however, more than made up for by the 
many disadvantages, as loss of metal by dusting and leaching, the 
consumption of fuel, slowness of roast, imperfect roasting, with 
consequent necessity of rehandling and reroasting, and obnoxious 
fumes. Therefore heap-roasting has not found much favor for 
matte. Where small quantities of matte are produced and where 
a calciner does not exist, heap-roasting is in place. 

A heap for roasting matte should not be very high, as the fire 
then becomes too hot and melts the central part of the heap ; on 
the other hand, if very low, the roast will be imperfect and the loss 
in silver and lead large. The sides of the heap are always covered 
with smalls, to prevent the fumes from passing off there. The 
height of a matte-heap varies from b\ to 1 feet ; the length and 
width have not much influence on the result. 

The roast-heap is usually built on the level of the feed-floor, as 
the fumes will not enter the smelter-building, and it is important 
not to carry the roasted matte any further than necessary. The 
loss incurred in transporting the raw matte to the roast-heap is not 
worth taking into account. The ground for a heap should be even, 
hard, and slightly elevated. Soft ground is levelled by a plough 
or a scraper ; hard ground is made even by filling ; the elevation 
of G inches is obtained by making a bed of coarse slag, covering it 
with fine slag, and rolling it, if possible. The first matte, being 
rich in iron, oxidizes and crumbles when exposed to the air for 
:i short time, so that the breaking-up into pieces of fist size does 
not cause much labor. In building the heap, a bed of light wood 
from '5 to 15 inches deep is made, leaving channels 10 inches wide, 
open every 6 feet, which are filled with kindling. Hard wood is 
not so good lor the bed, as it makes too hot a fire, which is liable 
to fuse the lower part of the heap. Sometimes a chimney is placed 
in the centre, consisting of four boards nailed together and long 
enough to reach 2 feet above the heap. It helps to start the heap 
uniformly. In building the heap, first coarse matte is piled around 
the chimney, and then distributed over the bed of wood. The 
coarse matte is assorted from the finer matte with a sluice-fork. 
When the heap has been built in the form of a pyramid the edges 
of which reach to within a {• foot of the border of the wood, the 
Bides aff covered with tine mat te and sometimes also the top. Fine 



256 



METALLURGY OF LEAD. 



matte is, however, kept on hand, to be used when the heap has 
been started, to make the top layer thicker, and to check the 
draught in parts of the heap that are becoming too hot. In Utah L 
the pyramidal heaps are made 24 by 18 feet at the base and 6 feet 
high ; they contain about 80 tons of matte, and burn from 30 to 40 
days. 

In the Hartz Mountains 2 matte-heaps are usually made 15 by 
23 feet on a bed of wood 6 inches deep ; they contain from 75 to 
100 tons and burn 30 days. Quite a saving of cord-wood has lately 
been effected by placing pieces of wood, split from sound dry 
logs about 9 inches in diameter, end to end in parallel rows, about 
12 inches apart, on the ground, and laying small sticks across 
them. On this bed are distributed in rows above the cord-wood 
large pieces of matte to the height of about 9 inches. Between 
these are piled faggots and soft coai. Looking at the heap from 
the top there will be seen alternate rows of coarse matte and fag- 
gots mixed with coal. On this bed the heap is built in the usual 
way. 

The well-roasted part of a matte-heap amounts to about 50 per 
cent., if the roasting has been carefully conducted and the sulphur 
has been reduced to about 6 per cent. Analyses 3 of raw and 
roasted matte from Clausthal are subjoined : 



Matte. 


Pb. 


Cu. 

4.620 
4.123 


Sb. 

0.267 

0.128 


Ag. 


Fe. 


Zn. 

2.110 
2.459 


Mn. 

0.385 
0.317 


Co. 

0.215 
0.239 


Ni. 

0.097 
0.111 


Si0 2 

0.510 
1.486 


CaO 

0.383 
0.336 


MgO. 

0.054 
0.061 


s. 


O. 


S0 3 . 
4.225 


Raw 

Roasted. 


10.665 
10.492 


0299 
0.0o27 


53 112 
52.411 


26.877 
0.613 


28.9663 



To roast one ton of matte in 80 or 100 ton heaps, about 0.3 
cords of wood are required and from 5 to 8 days' labor. 

2. Roasting in stalls has many advantages over roasting in 
heaps. There is less loss of metal by dusting and none at all by 
leaching; little fuel is needed; the piles can be smaller; the roast is 
quicker and more perfect, requiring less matte to be roasted a second 
time, and the fumes are carried off. The only drawback is the 
cost of plant, and this is more than made up, especially if the 
stalls are built of slag-brick, by the advantages. That stalls have 
not superseded heaps is because, if a plant has to be built, it is 

1 Terhune, Trans. A. I. 31. E., xvi., p. 23. 

2 Private notes, 1890. 

8 Arche, " Die Gewinnung der Metalle," Leipsic, 1888, p. 72. 



SMELTING IN THE BLAST-FURNACF. 



257 



considered that a long-hearth reverberatory furnace may as well 
finish the roasting by one short operation. The writer, however, 
does not agree with this view, and thinks that stalls might in many 
cases be preferable. 

The open stalls used at Pribram, 1 Figures 154-157, will serve as 
an illustration. The general arrangement is shown by Figures 156 
and 157. There is a double row of stalls placed back to back with 

Fig. 154 

ROAST STALLS FOR MATTE. 

FRONT VIEW OF STALLS, No. 4, 5, 6. 




PLAN AND HORIZONTAL SECTION. 




a flue between them, which is divided by a vertical wall into two 
parts, in older that each row of stalls may have an independent 
draught. The flue terminates in a chimney 82 feet high. The side- 
walls of a stall are built up solid; in the front wall on each side of 
the entrance are four fireplaces, from which the heap is kindled. 

The small Hues in the back -wall open into tWO chambers, w Inch eon- 



1 Zdrahal, Oestcrreichiscltes Jaltrbuclt, x.wix., p. U\, and Private 
notes, 1890. 



258 



METALLURGY OF LEAD. 



neet separately with the main flue, and can be closed off by sliding 
dampers. The advantage of the divided flue is that in treating 
small quantities of matte one side only of the stall is filled, and 
when the roast is finished, the clinkered and insufficiently-roasted 
parts can be placed directly on the heap which is being erected on 
the other side. Stalls 1, 2, and 3 receive small charges of 65 tons, 
and stalls 4, 5, and 6 large charges of 140 tons. The stalls are 
built entirely of slag-brick. They are made by tapping the blast- 
furnace slag from a slag-pot like that represented by Figures 134- 
139, and letting it run into a cast-iron mould. The upper surface 
is made smooth by placing a heavy cast-iron plate on the mould 
when filled. If the slag is not sufficiently acid, it cannot be used 
alone for this purpose, as it cracks while cooling. It must then be 



CROSS-SECTION. 




Vig. 157 
ROASTING-STALLS OF PRIBRAM. 

used to cement together clinkers, sharp-edged sand, pieces of brick, 
etc., placed in the mould. It is well to cover the mould with sand 
or ashes when filled, in order that the slag may cool slowly, as it is 
less liable to crack. Slag-bricks form an excellent building mate- 
rial for the stalls. They are laid with ordinary clay mortar; the 
walls are sufficiently thick not to require any binding. 
Four sizes are made: 



NO. 


LENGTH. 


WIDTH. 


HEIGHT. 


1. 


1734 inches. 


83^ inches. 


8kt inches 


2. 


n\i " 


43^ " 


834 " 


3. 


12* " 


12* " 


B)4 " 


4. 


12K " 


6^ " 


6^ " 



Bricks Nos. 1 and 2 are used for the side-walls; Nos. 3 and 
4 for the bottom of the stall. Nos. 2 and 4, having half the 
width of Nos. 1 and 3, correspond to split brick, and are used as 
binders. 



SMELTING IN THE BLAST-FURXACE. 



259 



Two improvements in construction may be mentioned. The 
first consists in leaving open two horizontal air-passages, one above 
the other, in each side-wall, with openings into the stall. One 
may be placed at half the height of the side-wall, terminating at 
half its depth; the other at three-quarters the height, ending at 
three-quarters the depth. The second is to leave openings in the 
front-wall. This introduction of air gives better control of the 
roast. 

In charging the stall, coarse matte is placed on a thin bed of 
wood arranged on the slag-bottom in the same way as when a heap 
is erected. Then follows a mixture of coarse and fine matte and 
the top is covered with sufficient fines to prevent the fumes from 
passing through the top, thus enabling them to be drawn off 
through the chimney. If only one-half of the stall is to be used, 
the open side is built up of coarse matte. The stalls are fired from 
the front with wood, and the fires kept going with bituminous coal 
and lignite until the heap is well ignited. 

The time required for roasting is : 





65 tons. 


HO tons. 


Unroasted lead matte 


48 days. 
43 << 
33 " 


86 days. 

80 «• 
70 " 


Once-roasted lead matte 


Twice-roasted lead matte 





The well-roasted matte contains from 8 to 10 per cent, sulphur, 
one-third of the sulphur being present as undecomposed sulphide. 
It makes no difference in this result whether the stall is entirely 



filled or only half filled. The change in 
insignificant 

The labor necessary per ton of matte is : 



weight after roasting ii 



For the first roast 

For t he Becond roast 

For the third roust 

For pulliiiLT down and wheeling to the 
feed-floor 



('>'> ions. 


ho tons. 


1 .SI hours. 

2 .26 " 
2.26 " 

8.62 " 


2.02 hours. 
2.44 •• 
2.44 " 



With four men working in the daytime only in ten-hour shifts, 
the number of days required to till and empty the stalls will be: 



260 



METALLURGY OF LEAD. 





65 tons. 


140 tons. 


Filling with raw matte for first roast. . 

Pulling down and filling for second 
roast 

Pulling down and filling for third roast. 

Pulling down and wheeling to feed- 
floor 


3 days. 
3*< " 

6 


7)^ days. 

8% " 
8% " 

13 


Handling 


wy 2 " 


37^ » 





T^he fuel required per ton of matte is : 





65 tons. 


140 tons. 


Wood, 
cords. 


Coal, 
pounds. 


Wood, 
cords. 


Coal, 
pounds. 


For first roast 


0.008 
0.008 
0.008 


68.2 
12.6 
71.6 


0.007 
0.007 
0.007 


66.0 

78.0 
84.4 


For second roas* - 


For third roast 




Average 


0.008 


70.8 


0.007 


76.1 



Analysis l of stall-roasted ore from the Colorado Smelting Co., 
Pueblo, Col.: 



Fe 9 a 


80.39 


Mn 2 O s 


0.93 


PbO 


7.91 


ZnO 


0.98 


CuO 


: 3.22 


As o 3 


0.86 


Sb 


trace 


S 


1.01 


SiOo 


3.21 


CaO 


3.22 


Aer 


0.0614 



3. Roasting in Kilns is practised only in connection with the 
manufacture of sulphuric acid. In the Hartz Mountains and in 
Saxony 2 shaft-furnaces are used for roasting. At the Lautenthai 
Works 3 six Freiberg shaft-furnaces, i each having four doors, are in 

1 Dewey, Bulletin No. 42, " United States National Museum," p. 52. 

2 Merbach, "Freiberg's Berg- u. Hiitten-Wesen," Freiberg, 1883, p. 251. 

3 Private notes, 1890. 

4 Drawings in Zeitschrift fur Berg-, Hutten-, und Salinen-Wesen in 
Preussen, xix., plate viii. 



SMELTING IX THE BLAST-FURXACE. 261 

"use. The inside dimensions are: height, 10 feet 6 inches; width, 4 
feet 2 inches; depth, 7 feet 5 inches. They hold about 300 cubic feet 
of ore. In twenty-four hours one furnace roasts in two charges 4,200 
pounds of a mixture of 4 parts raw first matte (4 per cent, copper, 
10 per cent, lead) and 3 parts half-roasted matte, or 5,400 pounds 
of a mixture of 5 parts raw second matte (12 per cent, copper, 7 
per cent, lead) and 4 half-roasted matte. The roasted matte 
contains from 7 to 8 per cent, sulphur, and the gases 5.3 volumes 
sulphurous acid. The six furnaces are attended by eight men in 
twenty-four hours. Sufficient heat is generated by roasting to 
have a Glover tower ; the sulphuric acid chambers have a capacity 
of 670,000 cubic feet, and terminate in a Gay-Lussac tower. In 
1879, when the writer was at Lautenthal, there was a loss in the 
manufacture of sulphuric acid, and it was carried on only to make 
the sulphurous acid innocuous ; now (1890) through the improve- 
ments in plant, acid is made with profit. Details about the 
management of furnaces and the manufacture of sulphuric acid are 
given in Lunge's excellent work. 1 

4. Roasting in Reverberatory Furnaces. — Of the different 
reverberatory furnaces only the long-bedded one-hearth calciner is 
in general use. Terhune 2 gives the results obtained at the Germania 
Works, Salt Lake City, in roasting matte in a Bruckner cylinder 
18 feet long and 7 feet in diameter. A charge of 8 tons is roasted 
in forty-eight hours down to 4 or 6 percent, sulphur, requiring one 
man in a 12-hour sliif t and consuming 20 per cent, of Pleasant 
Valley coal. Two cylinders, 22 by 7 feet, were added later, and all 
the three roasters are run by two men in 12 -hours shifts. From 
Planner's experiments, already mentioned, which show that the 
silver and lead in roasted matte is concentrated toward the surface, 
it can be Been that the loss in these metals by dusting must be 
great with revolving furnaces, and this is probably the cause that 
have not found more general application. 

Terhune 1 gives some interesting figures on a short experiment 
in roasting matte in a Stetefeldt furnace. 

The reverberatory furnaces used are very similar to those 

described for roasting ores (§ 50,6) except thai theyhave no fuse- 

and that the roasted matte is generally discharged through 

one or two openings in the hearth near the bridge into an iron oar 

1 " Sulphuric Acid and Alkali," London, 1891, vol. i. 
■ Trans a. I. M. /•:., svi., p. 19. 
8 Trans, a. I M. /•/.. avi., p, 83, 



262 METALLURGY OF LEAD. 

in the arched chamber below. Where ore-calciners are used, the 
matte is often roasted with the ore, if this is to be fritted or fused, 
furnishing the necessary iron flux. The details of manipulation, 
with a few slight exceptions, are the same as those given under the 
roasting of ore, and can be passed over. The hearth of the matte- 
calciner commonly used at present in Colorado has the same 
construction as that of the ore-calciner. It is 60 by 14 feet, and 
is divided into four separate hearths, each having two working- 
doors on either side. It receives a charge of 5,400 pounds, crushed 
through a 4-inch sieve, and lies three or more inches deep in the 
furnace. In twenty-four hours three charges are drawn (generally 
into slag-pots) by four men, and two tons of coal are consumed. 
Pyrite often replaces part of the matte in the charge. 

The details of a matte-roasting furnace, which is also used for 
ore that is not to be fused, are given in Figures 158-163. It is one 
of the furnaces contained in the roaster building shown in Figure 
82, and is chosen because it differs considerably in its constructive 
details from the type generally accepted in Colorado as standard. 
The hearth (14 by 40 feet) has about the same width as the 
Colorado furnace, but is considerably shorter. The reason is that,, 
as the charge is not to be fused, the temperature is kept compara- 
tively low at the bridge, which permits a 40-foot hearth to utilize 
the heat about as well as a 60-foot one. Sj^ecial attention is called 
to the vaulted arches which support the hearth, the absence of off- 
sets (the hearth having a gentle rise from bridge to flue), the slope 
from the centre to the doors, the discharge-openings for roasted 
ore, and the damper in the flue. The distance between the work- 
ing-doors is less than is usually found ; the width of the fireplace 
(21 inches, Figure 162) proving insufficient, it was increased to 36 
inches (Figure 158), and to supply some of the room required for 
this, the fire-bridge (27 inches, Figure 163) was made narrower (22^ 
inches, Figure 158) and received air-flues. 

§ 77. Smelting of Roasted Matte. — The first matte ob- 
tained in ore-smelting (containing about 4 per cent, copper) is, 
after being roasted, always added as iron flux to the ore-charge in 
the blast-furnace. The resulting second matte, containing from 10 
to 13 per cent, copper, is concentrated in the blast-furnace, in 
the reverberatory-furnace, or in the two successively. 

1. In the Hartz Mountains the matte is concentrated exclusively 
in the blast-furnace to black copper (76 per cent, copper) and an en- 
riched copper matte (67 per cent, copper). As many as four sep- 
arate heap-roastings, each followed by a reducing smelting in the 






J 






A 







.ear .on ioa b ,oo hoito33 










.ear .omen 






Q 






















fi 



■ 












/ 


^Vi<^^>- 









I I Tw 



-i — t-- 



JECTMN I" FOFF 







, fi 


mig. 161 




;.i 




*rt* 




-J 

^~ 1 L 

"I 




— r-rhr 

















Bf 



SECllO.'i ■„!,,(■ A OF FIG. 1.60, ANDB /..rAOFF 





- 




' 




— 6V — 


„ 






J— 









^Jl] 







:i oir 



EM g . 150 
FOUNDATION PLAN OF BRICK WORK REV'Y ROASTING FURNACE. 



i ~ r 



* A 











\x/ J f v 


1 


6 

3 HOIT032 ' 






. 









SMELTING IN THE BLAST-FURNACE. 



263 



blast-furnace, are carried on to obtain the final products that are 
sold to copper works. 

The roasted matte is smelted with 147 per cent, slag and 15 per 
cent, coke in a low blast-furnace having a crucible that is partly in- 
ternal and partly external, or only external. 

One hundred parts of roasted second matte give 13.06 parts of 
impure base bullion and 35.42 parts of third matte. From this 
on no more lead can be extracted. The rate at which the matte 
decreases in quantity as its percentage of copper increases is shown 
by the following tables: 

AMOUNT OF MATTE. 

First matte 60.27 

Second " 21.00 

Third " 10.96 

Fourth 4 ' 4.57 

Fifth " 3.20 

Total 100 . 00 

ANALYSES OF MATTE. 1 



Pb 

Ak 
Cu 
Fe. 
Zn 

S... 



Mattel. Matte II. Matte III. Matte IV. Matte V 



13.47 
0.035 
5.73 

48.18 
1.71 

25.25 



8.33 

0.040 

12.86 

44.60 

3.32 

20.42 



10.02 
0.065 

27.76 

31.25 
2.99 

21.92 



9.06 

0.075 

42.30 

19.98 

1.59 

17.89 



3.88 

0.035 

59.86 

12.24 

0.75 

20.78 



ANALYSES OF SLAGS. : 



Si0 2 . . 
BaS0 4 
Pb.... 
Cu . . . . 
Ag.... 
Sb . . . . 
FeO . . 
Al 2 O a . 



MnO . . 
ZnO... 
Ni+Co 
CaO . . 
MgO... 
BaO .. 



Na 2 0. 
S 



23.73 

1.68 
0.58 
0.002 
0.14 
48.16 
8.12 
0.41 
0.60 
8 11 
0.09 
3.60 
0.32 
41 
0.42 
0.36 
3.95 



25.30 

0.36 

1.53 

0.63 

0.0019 

0.05 
51.99 

3.86 

0.56 

0.71 

6.69 

0.08 

4.61 

0.40 

0.47 
33 
3.49 



27.43 

1.06 

1.42 

56 

0.0013 

0.14 
49,67 

5.28 

1.31 

0.77 

6.38 

0.12 

4.16 

0.52 

0.64 
0.33 
2.39 



17.43 
trace 

1.47 

1.90 

0.0015 

0.09 
62.14 

5.87 

0.43 

0.49 

4.30 

0.27 

2.00 

0.39 

0.57 
0.33 

2.64 



18.97 

none 

0.92 

1.58 

0.20 

61.84 
6.47 
0.47 
0.50 
2.90 
0.54 
2 27 
0.47 

0.77 
0.53 

1.48 



Private notes, 1890. 



264 METALLURGY OF LEAD. 

From the above analyses can be seen that the roasting of the 
second matte is imperfect; the results after smelting show an in- 
crease of 15 per cent, copper instead of 30 per cent., as might have 
been expected. Thus the fourth matte is only as high as the third 
should have been. While the amount of iron in matte V. has been 
somewhat reduced in comparison with that of matte IV., it is still 
very high, and can be removed only by smelting in the reverbera- 
tor)- furnace. It would seem, therefore, advisable to stop the con- 
centration in the blast-furnace at 40 or 50 per cent, copper, and 
to treat the resulting matte in the reverberatory furnace. 

2. The general practice of American smelters is to add the 
roasted first matte as iron flux to the ore-charge and to treat the 
resulting second matte in a separate furnace. This has the same 
general construction as the lead-furnace, with the exception that it 
has no crucible to speak of, the distance between centre of tuyere 
and slag-tap being ten or twelve inches. Often the crucible of a 
lead-furnace has been simply filled with well-beaten brasque and 
is thus changed into a matte-furnace, but special furnaces with a 
drop bottom, such as is common in copper-smelting, are often 
used. 

The dimensions at the tuyere level vary from 36 by 60 inches 
to 33 by 100 inches, according to the required capacity, the latter 
smelting 100 tons charge in 24 hours. In separating matte from 
slag the practice varies. 

At some works the tapping-jacket has two tap-holes, the lower 
one for lead and matte, the upper one for slag. The pot containing 
lead and matte with some slag is allowed to cool on the dump. 
When cold the button of lead is separated from the matte, and this 
from the slag, the latter being smelted over again. The slag from 
the upper tap-hole is also allowed to cool on the dump, and is 
examined for shots of matte before it is discarded. 

At other works the entire contents of the furnace are tapped into 
an overflow-pot. This is to collect the lead, while matte and slag 
overflow into a catch-pot having a tap-hole. All the slag remaining 
in the second pot is smelted over again. The overflow-pot has a 
tap-hole placed on the side, as near the bottom as possible. Every 
six or twelve hours, according to the grade of matte that is being 
smelted, this pot is wheeled over a cast-iron basin let into the ground, 
and the lead tapped. As soon as matte apjjears the tap-hole is 
olosed with a clay plug. 

A third method of obtaining a good separation of slag and 



MELTING IN THE BLAST-FURNACE. 265 

matte is to have the bottom of the furnace inclined. Thus the 
distance between the centre of the tuyere and the slag-tap is 10 
inches, and this is 8 inches higher than the matte-tap placed at the 
opposite end of the furnace. An excellent separation of matte and 
slag is effected. 

The roasted matte is usually smelted with foul slag from the 
ore- furnace and with coj^per ore low in silver, if this be available. 
The sla^s made are similar to ore-furnace slas^s, w T ith this difference, 
that they usually run lower in lime and are not quite so siliceous. 
Two special matte-slags may be quoted: 

Si0 2 . FeO. CaO. 

32 36 14 

32 40 16 

The reason that the roasted matte is not smelted with dry sil- 
ver ores is that only 93 per cent, of silver contained in the matte 
is paid for by copper refiners. The tendency is therefore not to en- 
rich, but rather to reduce the silver contents of the matte. The 
first matte obtained in the blast-furnace runs from 24 to 30 per 
cent, copper; when this is smelted again, after having been roasted 
in a calcining furnace, its copper contents range from 40 to 60 per 
cent, when it goes to copper works. 

3. The concentration of 12 per cent, matte in the reverberatory 
furnace alone is not practised. At Freiberg 1 it is brought for- 
ward to about 50 per cent, copper by two smeltings in the blast- 
furnace. Analyses of matte and slag by Schertel 2 are subjoined: 



MATTE. 


SLAG. 


Pb 


16.86 

0.29 


SiO, 


32.71 


Ag- 


PbO 

CuO 


8.16 


Cu 


. . . . 55.43 


. . . 1.40 


Sb 


. , . . , 0.22 


FeO 


. . . 38.01 


Fe 


3.50 


(Ni, Co)0 





Ni, Co 


0.41 


ZqO 


. . . 9.14 


Zn 


3.44 


A1 2 3 


. . . 5.30 


S 


20.12 


BaO 


. . . 2.88 






CaO 


. . - 0.59 



MgO — 

S 0.' 



The matte is then roasted in a calcining-furnace and smelted in 

1 Grand, Annates des mines, 1875, vii., p. 314 ; Capacci, Revue uni- 
verselle des mines, 1881, ix , p. 269. 

8 Berg- u. Hilttenm. Zeitung, 1888, p. 442. 



266 METALLURGY OF LEAD. 

a reverberatory furnace with barite and quartz as fluxes. The 
hearth of the furnace is 13 feet long, 4 feet 2 inches wide at the 
bridge, 8 feet at the middle, and 1 foot 2 inches at the flue; the 
grate, 4 feet 3 inches by 3 feet 4 inches, is 1 foot 3 inches below 
the top of the bridge, and this 1 foot 3 inches below the roof. The 
furnace has the usual two doors, one at the side and one beneath 
the flue. The bottom, which is built on an iron plate, consists 
first of a full course of fire-brick on which is rammed firmly a mixt- 
ure of fire-clay and quartz, giving it the usual disc-like form, 
with the lowest point at the tap-hole. The thinnest part of this 
layer is 3 inches thick. The working-bottom, which follows, con- 
sists of 5,500 pounds of an intimate mixture of 5 parts of quartz 
and 1 part of matte slag. It is melted down, after being heated 
and patted into shape, in twelve hours, and lasts from 15 to 18 
months. Five charges, each weighing 4,290 pounds, and consisting 

of 

2,640 lbs. roasted matte, 
440 " raw matte, 
660 " barite, 
550 " quartz, 

4,290 lbs. 

are treated in this furnace in twenty-four hours, 4J tons of bitu- 
minous coal being consumed. From these five charges 8,250 pounds 
of matte is produced. The composition of matte and slag as ana- 
lyzed by Schertel l is given below. 



MATTE. 




SLAG. 




Pb 


... 4.85 


Si0 2 


. . 30.60 


Ag 


. .. 0.31 


PbO 


. . 14.51 


Cu 


... 73.95 


CuO 


. . 4.40 


Bi 


. . . 0.02 


FeO 


.. 8.10 


Sb 


0.06 


(NiCO)O 


.. 0.38 


As 


. .. 0.18 


ZnO 


.. 1.95 


Fe 


. .. 0.13 


A1 2 3 


. . 1.80 


NiCo 


... 0.21 


BaO 


. . 28.63 


Zn 





CaO 


. . 7.70 


S 


. . . 18.98 


MgO 


. . 0.72 






S s 


. . 0.18 



In order to obtain matte running so low in iron, it is necessary 
to produce some bottoms. These go back to the next reverberatory 
charge or are added to the last blast-furnace charge to be again 

1 Log. cit. 



SMELTING IN THE BLAST-FURNACE. 267 

converted into matte. The matte forms the raw material for the 
manufacture of blue vitriol. 

This Freiberg method of using barite in the final concentration 
in the reverberator}' furnace will hardly recommend itself, as no- 
body will introduce baryta into a blast-furnace slag if it can be 
helped. By substituting gypsum, which, as shown by Schweder 
(§ 54, g), has the same effect, slags low in copper might be ob- 
tained, and the resulting slag, containing lime instead of baryta, 
would be welcome for the furnace charge. Thus it would be nec- 
essary for only a small amount of the copper from the concen- 
trated matte to pass again through the lead blast-furnace and 
undergo the tedious nrocess of matte concentration. 

4. Treatment of Argentiferous Copper Matte containing Lead 
and Iron. The concentrated matte produced at lead works is 
usually shipped, as before stated, to copper works, and there 
treated by one of two methods: the first is to convert the matte 
into metallic copper and refine it, after which it is desilverized by 
electrolysis, vitriolization, or a modified Augustin process ; the 
second is to dead-roast the matte, if it is sufficiently free from iron 
as at Freiberg (see ante, a), and then treat it with sulphuric acid 
to dissolve the copper, which is sold as blue vitriol, the residue 
going to a lead blast-furnace. 

The aim at lead works for some time has been to work up the 
matte there. This is done in two ways — by the Crooke process and 
the Hunt & Douglas process, No. 2. 

a. The Crooke Process. — Several years ago John J. Crooke in- 
vented 1 a process to desilverize copper matte and produce metallic 
copper. It is now in use at the works of the Pueblo Smelting and 
Refining Co., Pueblo, Col., but, as details are not published, only a 
short outline can be given. 

The desilverization is based on the property lead possesses of 
extracting silver and gold from copper matte. If lead is kept at a 
red heat and crushed matte spread over it and allowed to remain 
in contact, from To to 80 per cent, of the silver and all the gold 
will be extracted. The lead, however, takes up copper, and if this 
goes too far it loses the power of desilverizing. Also, if it remains 
in contact with the matte beyond a certain time, the two are liable 
to combine into one mass. To prevent this, Crooke introduces into 
the lead bars of the best soft Swedish iron, placing them horizon- 

1 Patents No. 308,031, November 11, 1884 ; No, 354,182, December 14, 

1886. 



268 METALLURGY OF LEAD. 

tally across the bottom of the reverberatory furnace and fastening 
them there. To extract the remainder of the silver, fresh lead is 
required. 

In actual practice four reverberatory furnaces are used, each 
holding 25 tons of lead, arranged in terrace form, in order that the 
lead of the upper one can be discharged into the next one below. 
Into the highest furnace is charged lead free from or low in silver. 
This will be gradually enriched in its passage through the furnaces, 
until taken out at last as rich bullion, containing 2 or 3 per cent, 
copper and a little arsenic and antimony. The rich matte, 
charged into the lowest furnace, will be gradually desilverized in 
its passage upward over the four baths of lead, and will be drawn 
from the highest furnace as granular matte practically free from 
silver. About twenty-five tons of matte are thus passed in 
charges of one and a half tons each in twenty-four hours through 
each of the four reverberatory furnaces, taking about one and 
a quarter hours for the three operations : charging, treating, 
and withdrawing. 

The second part of the process, the conversion of desilverized 
copper matte into metallic copper, requires four operations, which 
are carried on in four separate furnaces. The matte is first roasted 
in a calcining-furnace to reduce the sulphur contents down to 
13 per cent. It then passes to a reverberatory furnace having 
a tuyere on either side of the bridge. Here it is subjected to a 
higher temperature, which is, however, kept below the melting- 
point of copper. By the action of heat and blast and by constant 
rabbling, the impurities of the matte will be oxidized, and the cop- 
per converted into moss-copper. When this has formed, the 
charge is transferred to a blister-furnace and melted down quickly, 
with an addition of quartz to slag the lead and any iron that may 
be present. The slags containing lead, iron, arsenic, and antimony 
are drawn ; then follow coppery slags, and when the charge comes 
to work, the blister-copper, assaying 99 per cent, copper and over, is 
tapped into sand-moulds, and is then ready for the refining-f urnace. 

b. The Hunt an d Uovglas Process. JVo. 2. 1 This process, in- 

1 United States Patent, No. 227,902, May 25, 1880 ; Hunt, Trans. A. T. 
M. E., x., p. 12; xvi., p. 80 ; Douglas, '• Mineral Resources of the United 
States," 1882, p. 279 ; Howe, " Production of Gold and Silver in the United 
States, Report of the Director of the Mint," 1883, p. 798 ; Hunt and Doug- 
las, Engineering and Mining Journal, October 3, 1885 ; Fulton, Ibid., De- 
cember 12, 1885 ; Howe, Ibid., December 19, 1885. 



SMELTING IN THE BLAST-FURNACE. 269 

vented over ten years ago, has lately been introduced at the lead- 
silver works of the Consolidated Kansas City Smelting and Refin- 
ing Company, Argentine, Kan. It has not been running long 
enough to furnish final figures, although so far it appears to be do- 
ing satisfactory work. A general outline must therefore suffice for 
the present. The process consists in : 

1. Dissolving the copper from roasted matte with dilute sul- 
phuric acid containing some chloride, the residue, which goes to 
the blast-furnace, being iron, lead, and silver. 

2. Chlorodizing the cupric sulphate with a solution of ferrous 
or calcium chloride which contains a slight excess of chloride over 
what is required to form cuprous chloride. If calcium chloride is 
used, nearly all the calcium will be precipitated as calcium sul- 
phate, and have to be separated by an additional operation. 

3. Reducing the cupric chloride formed to cuprous chloride, and 
precipitating this by pumping hot concentrated sulphurous acid, 
generated by burning pyrite in the Douglas " central. flue cylinder 
calciner," ' into the solution, an equivalent amount or more of 
sulphuric acid is formed. This is freed from sulphurous acid by 
injecting hot air, which also completely oxidizes any metallic salts 
held in solution. It is then ready to be used as solvent for roasted 
matte when it has cooled. Any iron, zinc, nickel, or cobalt salts 
gradually accumulating in the solvent have to be removed period- 
ically by crystallization. 

4. Converting the cuprous chloride by means of metallic iron 
into metallic copper, or by milk of lime into cuprous oxide, to be 
smelted for marketable copper. The ferrous or calcium chloride 
formed by the reaction serves as chlorodizing agent for the cupric 
sulphate of the next charge. 

§ 78. Nickel and Cobalt Matte. — The behavior of nickel 
and cobalt in lead matte shows some points of interest. The non- 
argentiferous ores of Southeast Missouri (§ 19) carry small amounts 
of nickel and cobalt, which are concentrated in a blast-furnace 
with syphon-tap in a matte. At Mine La Motte 2 the first matte car- 
ries from 3 to 3.5 per cent, nickel-cobalt, from 0.5 to 1 per cent, cop- 
per, and from 20 to 25 per cent. lead. "When this is roasted in a cal- 
cining furnace, there must remain, when it comes to be used as. iron 
flux, from 5 to 6 per cent, sulphur in the matte, if the loss of nickel 

1 Douglas, Engineering and Mining Journal, August 14, 1886 ; Aaron, 
Ibid., September 11, 1886. 

2 Neiil, Trans. A. I. M. E., xiii., p. 634. 



270 METALLURGY OF LEAD. 

and cobalt by slagging is to be avoided. In the same way the second 
matte (5-6 Ni-Co; 20-30 Pb; 1-2 Cu), when roasted, requires from 
7 to 9 per cent, sulphur to reduce the loss by slagging ; it is then 
smelted with siliceous matter to a third matte (12-17 Ni-Co ; 35-40 
Pb ; 3-5 Cu), the slags nevertheless assaying L25 per cent, nickel- 
cobalt and from 2 to 2.5 j:>er cent. lead. Attempts at further 
concentration in a reverberatory furnace proved unsuccessful, the 
slags carrying 5 per cent, nickel-cobalt and 5 per cent. lead. This 
illustrates the difficulty of concentrating nickel-cobalt in a matte 
when lead is present, which apparently drives it into the slag. The 
nickel can then be recovered only in part and the cobalt not at all. 
The way that remains is to introduce arsenic and form a speise. 
Neill used this method successfully. The experimental furnace he 
built was 30 inches wide at the tuyeres, and 48 inches at the 
charging-door ; the height from tuyeres to charging-door was 
6 feet 6 inches ; the crucible, 18 inches deep, was partly internal, 
partly external, and had a syphon-tap. He obtained a speise with 
22.53 per cent, nickel-cobalt (^ of which was cobalt), 6.4 per cent, 
lead, and 4.25 per cent, copper. The matte that formed contained 
3.25 per cent, nickel- cobalt, 8 per cent, lead, and 7 per cent, copper 
to be retreated ; the slag from 0.16 to 0.32 per cent, nickel-cobalt 
and from 0.6 to 0.8 per cent. lead. 

§ 79. Analytical Determinations. — The bodies to be de- 
termined by assay in raw or roasted matte are silver and gold 
{§ 34), lead, copper, iron, and zinc (§ 81), and sulphur. 

1. Lead. — The usual dry assay for lead (§ 34) gives un- 
satisfactory results, especially in raw matte. Only from 40 to 50 
per cent, of the lead present is collected in the button. If the 
sulphur and metals other than lead are removed, the results ob- 
tained with the usual crucible-assay are accurate. To do this, 
5 grammes of matte are dissolved in a small flask with nitro- 
hydrochloric acid ; dilute sulphuric acid is added, and the whole 
evaporated to dryness and heated till sulphuric acid fumes are 
given off ; when cool, hot water is added, the solution boiled, 
filtered, washed, the filter incinerated on a scorifier in the muffle, 
and the lead sulphate with any insoluble residue reduced in the 
crucible in the usual way. 

Williams l recommends a wet method, devised by von Schulz 
and Low, of Denver, Col., which is suitable for ores as well as by- 
products. It is as follows. Dissolve 1 gr. ore in 10 cc. pure con- 
1 Engineering and Mining Journal, June 18, 1892. 



SMELTING IX THE BLAST-FURNACE. 271 

oentrated nitric acid, and 10 cc. pure concentrated sulphuric acid, 
heat till fumes of sulphuric acid are given off copiously, cool, add 
10 cc. dilute sulphuric acid (1 acid : 9 water), then 2 grs. Rochelle 
salt, and when this is dissolved, 40 cc. water ; heat to a boil, allow 
to settle, filter and wash with dilute sulphuric acid. Dissolve the 
lead sulphate from the filter with a boiling concentrated solution 
of ammonium chloride held in a wash-bottle, and collect the filtrate 
in a flask in which have been placed three bits of aluminum foil 
(size | by If inches, thickness -^ inch.) Wash with the solvent, 
using little of it, boil the filtrate for five minutes, give the flask 
a rotary motion to collect the precipitated metallic lead, fill with 
cold water, invert in a casserole, discharge the contents, remove 
the aluminum foil, rubbing off any adhering lead with the fingers, 
wash with water by decantation and finally with alcohol. Collect 
and press the lead together with an agate pestle, warm gently to 
avoid oxidation, transfer to the scale-pan and weigh. Assays 
made on ores by this method give results that are three per cent, 
higher than those obtained with the usual fire assays. The time 
required is less than forty minutes. 

2. Copper. — The copper in the matte is determined by titrat- 
ing with potassium cyanide or by electrolysis. The former method 
is universally used in estimating the copper in matte that is to be 
further concentrated; the latter is more common with enriched 
matte that is to be sold, although the cyanide method also finds 
much favor. 

a. The Cyanide- Assay} — From 2 to 5 grammes of matte are 
dissolved in nitro-hyrochloric acid and sulphuric acid, evaporated 
and filtered as above, the filtrate being caught in a porcelain cas- 
serole. The copper is precipitated with granulated zinc, care being 
taken that no zinc remains with the copper by dissolving the last 
particles with dilute sulphuric acid. The solution is poured off 
and the copper washed by decanting and transferred to a beaker. 
From 8 to 16 cc. nitric acid (1.2 spec, gr.), diluted with an equal 
amount of water, is added, the solution is slightly warmed, then 

1 Beringer, CJiemical Neics, xlviii., p. Ill ; Berg- u. Hutt. Ztg., 1884, 
p. 200 ; Brugman, Engineering and Mining Journal, May 18, 1889 ; Fes- 
senden, Ibid., May 31, 1890 ; Low, Proceedings Col. Scientific Society, vol. 
i., pp. 17, 69 ; Berg- u. Hilttenm. Ztg., 1886, p. 53 ; Peters, "Modern Cop- 
per Smelting," New York, 1891, p. 33 ; Torrey and Eaton, Engineering and 
Mining Journal, May 9, June 6, 27, 1885 ; Wendt, School of Mines Quar- 
terly, vii., p. 222; Eustis, Trans. A. I. M. E., xi., p. 120. 



272 METALLURGY OF LEAD. 

cooled, made amraoniacal with 10 oc. of dilute ammonia (1 am- 
monia : 2 water), and titrated with potassium cyanide (20 grs. potas- 
sium cyanide to 1 liter water). To obtain satisfactory results with 
the assay the usual conditions of titration must be complied with, 
viz., that a given volume of solution (200 cc.) shall contain the same 
amount of acid and ammonia and the same temperature prevail as 
in standardizing the cyanide solution with pure copper. As regards 
the finishing point of the titration, it is customary to stop at a cer- 
tain shade of pink, which each assayer fixes for himself. A more 
satisfactory way is to add cyanide until the solution becomes de- 
colorized. This will be when a tube filled with the solution shows 
the same colorlessness as one filled with distilled water, when they 
are held together perpendicularly over a sheet of white paper and 
examined from above. Should the tube with the copper solution still 
show a slight shade of pink, it is returned to the main solution, 
a drop or two of cyanide is added, and the comparison repeated. 
In this way the slightest differences in color are easily detected. 

In the decanted solution the iron can be titrated with potassium 
permanganate (§ 81). The residue from the solution can be weighed 
after the lead sulphate has been dissolved by ammonium acetate or 
citrate. 

b. The Battery -Assay}— The filtrate from the lead sulphate is 
neutralized with ammonia, 3 cc. of concentrated nitric acid are 
added, the liquid containing not more than 100 cc. A current that 
gives from 90 to 100 cc. of hydrogen and oxygen in 30 minutes 
from dilute sulphuric acid (1 : 12) will finish the precipitation in 
from 8 to 12 hours, the time required varying according to the 
amount of other salts present. The deposition can be hastened 
by removing the iron and other salts, i.e., by precipitating the 
copper with zinc as in the cyanide-assay, and redissolving in ni- 
tric acid. To ascertain if all the copper has been precipitated, 
the solution is stirred and a little water added. If no copper is 
precipitated on the j)latinum dish or the platinum cone or cylinder 
in the beaker, the precipitation is complete. The other test is 
with hydrogen sulphide. If bismuth, arsenic, and silver are present, 
they are determined separately and deducted from the metal depos- 

1 Low, Col. Scientific Society, i., pp. 17, 69 ; Berg- u. Huttenm. Ztg., 
1886, p. 53 ; Torrey and Eaton, Engineering and Mining Journal, May 9, 
June 6, 27, 1885; Peters, "Modern Copper Smelting," New York, 1891, 
p. 39 ; Eustis, Trans. A. I. M. E., xi., p. 120 ; Glenn, Trans. A. I. M. E., 
xvii., p. 406. 



SMELTING IN THE BLAST-FURNACE. 273 

ited. The silver may be completely deposited with the copper 
if a few drops of tartaric acid be added to the solution of cupric 
nitrate. 

3. Sulphur. — Two methods are used at lead works to determine 
the sulphur. Both aim to convert it into sulphuric acid, which is 
precipitated by barium chloride, and the resulting barium sulphate 
weighed. Before precipitating, both solutions should be brought 
almost to a boil and all the precipitant be added at once. In this 
way the barium sulphate will settle quickly and the filtrate be 
clear. 

1 BaS0 4 contains 0.3433 S0 3 or 0.1373 S. 

a. Fuse 1 gr. matte (ore, slag) with 10 grs. of a mixture of 2 po- 
tassium nitrate and 3 sodium carbonate in a porcelain crucible. 
Heating to quiet fusion is not necessary to oxidize all the sulphur. 
Boil the fused mass with water, introduce carbonic acid to precip- 
itate any lead dissolved by the alkaline solution, filter, add hydro- 
chloric acid, evaporate to dryness, take up with hydrochloric acid, 
filter the silica, boil, precipitate with boiling barium chloride. 
The evaporation to dryness is avoided by the second method. 

b. Fuse 1 1 gr. matte (ore, slag) with 25 grs. caustic potash in 
a silver or gold-lined platinum crucible for 20 minutes; cool, dis- 
solve in water, filter, add 30 cc. bromine water, then hydrochloric 
acid; boil, filter, add barium chloride, etc. 

§ 80. Slag. — The composition (§ 53) and the disposal (§ 65) of 
lead slags have already been fully discussed. 

If slags are too rich in silver to be thrown away, and re-smelt- 
ing is of no avail (i.e., when the ore contains considerable amounts 
of zinc), they can be desilverized by smelting with copper-bearing 
pyritic ore or matte. Keller 2 smelted slag, estimated to run 5.3 
ounces silver per ton, with 13 per cent, pyritic ore containing 10 per 
cent, copper and 11 ounces silver per ton in an oblong blast-furnace 
36 by 80 inches. He put through in twenty-four hours 113 tons of 
charge with 10.9 per cent, coke, a little more than half the amount 
required to smelt ore-charges. There resulted 5.3 per cent, matte 
containing 20 per cent, copper and 92.7 ounces silver per ton, show- 
ing a saving of 80 per cent, of the silver contained in the original 
slag. 

1 Fahlberg-Iles, "Mineral Eesources of the United States," 1883-84, 
p. 450. 

2 Trans. A. I. M. E., xx. 



274 MET ALL URGY OF LEAD. 

§ 81. Analytical Determinations. 1 — Slags are assayed for sil- 
ver, gold, and lead (§ 34), silica, iron, manganese, lime, baryta, mag- 
nesia, alumina, and zinc. The aim in making the determinations must 
be quickness, even if it be somewhat at the expense of accuracy. 
If a furnace is not working well, a satisfactory correction in the 
charge can only be made if the main constituents of the slag are 
known, and as it takes from six to eight hours for a new charge to 
have any effect on the outgoing slag, it is easy to see that the 
quicker the results of the analysis can be obtained the better for 
the furnace. For this reason it has been necessary to adapt the 
usual analytical methods in such a way that quick results may be 
obtained. The cardinal rule is to use as little of the reagents and 
wash water as possible. 

An important feature is the complete decomposing of a slag 
without having resort to fusion with alkalies. All lead slags 
containing 35 and less per cent, silica are decomposed by hydro- 
chloric acid if the slag sample is glassy and sufficiently pulver- 
ized. That glassy slags are more easily decomposed than crypto- 
crystalline can be explained by the same reasons Le Chatelier 2 gave 
for the fact that an iron blast-furnace slag, when chilled suddenly, 
shows the properties of a natural cement. If a slag is suddenly 
chilled and becomes vitreous, it still retains the heat of crystalliza- 
tion, w T hich would be set free if the slag cooled slowly in the pot 
and became crystalline. The components are therefore in a less 
stable condition, and are easily decomposed by the acid. In taking 
the sample from a freshly-drawn pot of slag the hardened surface 
is perforated, and a clean steel bar inserted 3 inches into the liquid 
slag and quickly chilled by plunging into cold water. Sometimes 
slag is dipped out from the pot w r ith a clean cold iron ladle, 
poured out again after a minute or two, and the thin shell of 
glassy slag adhering slightly to the ladle taken as sample. The 
former method, however, is preferable. The sample is crushed, 
cut down, and ground so as to pass a 120-mesh sieve. Care should 
be taken that the entire final sample should pass through the 
sieve. The part to be weighed out should be ground for a 
minute in an agate mortar before placing it on the tared watch- 
glass of the balance. This grinding is sometimes omitted, but it 
may save a great deal of trouble later on. 

1 lies, " Mineral Resources of the United States," 1883-84, p. 449 ; Koh- 
ler, Engineering and Mining Journal, August 25, 1885. 

2 Annales des mines, 1889, xvi., p. 165. 



SMELTING IN THE BLAST-FURNACE. 275 

1. Silica. — Add to 0.5 grs. slag in a 3^-inch porcelain casserole 
a "pinch " of finely-ground potassium chlorate, mix with a glass rod, 
add 2 or 3 drops of water, rub to a paste, add 5 cc. cone, hydro- 
chloric acid, heat, add a few drops of nitric acid, evaporate to dry- 
ness over a lamp or on a sand-bath, stirring constantly. To prevent 
spattering, spread the mass over the sides of the casserole when 
nearly dry, and rub the residue down from the sides into a small 
heap. If heated too much, some iron becomes insoluble. Add a 
few drops of hydrochloric acid, evaporate again, then add 15 cc. 
cone, hydrochloric acid, heat, dilute with a little water, boil, filter, 
wash three times, dissolve any lead with ammonium acetate, wash 
three times, and discard the filtrate. Transfer the wet filter to a 
platinum crucible, ignite in the muffle, and weigh. This gives the 
silica and perhaps the insoluble silicate. To test for the latter, 
add twice hydrofluoric acid, evaporate over a lamp under a hood, 
ignite, and weigh again. The difference in the two weights is 
silica; the difference between the last weight and the weight of 
the crucible is the insoluble silicate. 

Another way of eliminating lead sulphate is given by lies. 1 
When the hydrochloric acid solution has been evaporated to dryness, 
take up with hydrochloric acid, dilute, warm, add metallic zinc, 
which reduces to metal any lead, silver, and copper, and to a fer- 
rous salt the ferric chloride. Filter and titrate the iron in the fil- 
trate (§ 81, 2), dissolve the reduced metals, leaving as residue silica 
(and silicate). 

If lead sulphate and undecomposed silicate need not be consid- 
ered, transfer the wet filter to an annealing-cup, ignite in the 
muffle, empty the silica on a tared watch-glass, and weigh. 

In an ore containing barite and a silicate that is not decomposed 
by acid, the insoluble residue will be treated as follows. Fuse with 
4 parts of sodium carbonate, giving a cover of sodium carbonate, 
disintegrate in hot water, add hydrochloric acid, evaporate to 
dryness, take up with hydrochloric acid, dilute, boil, filter, wash, 
ignite, and weigh silica plus barium sulphate. Volatilize the silica 
as silicon fluoride with hydrofluoric acid and weigh again; the 
difference is barium sulphate. Or, fuse again with four parts sodi- 
um carbonate, disintegrate in water only, filter off the sodium sili- 
cate, wash, dissolve the barium carbonate from the filter with 

1 School of Mines Quarterly, iii., p. 223. 



276 METALLURGY OF LEAD. 

hydrochloric acid, wash, precipitate hot with hot dilute sulphuric 
acid, and weigh the barium sulphate. 

•Another method is given by Furman. 1 

lies 2 recommends carrying on the fusion in a silver crucible 
with caustic potash. Fuse 40 or 50 grs. caustic potash in a silver 
crucible (holding 70 cc. water) to quiet fusion, cool, add 1 gr. pul- 
verized substance, heat for 30 minutes. The caustic alkali acts in 
the same way as the carbonate. The advantages of this method 
are quickness of fusion at a low temperature and cheapness of cru- 
cible. 

2. Iron. — The iron is titrated either with potassium bichromate 
(5 grs. of the powdered and dried salt to 1 liter of water), using 
potassium ferrocyanide (0.1 gr. in 100 cc. water) as an indicator 
on a dry white porcelain plate, or with potassium permanganate 
(5.686 grs. to 1 liter of water). Both solutions are standardized with 
0.2 grs. iron wire (weight X 0.997 = Fe). 

a. Bichromate Method. — Decompose the slag as in determining 
silica, but do not filter or use the filtrate of the silica if this has 
been determined. Add to the hot solution in the beaker stannous 
chloride (20 grs. stannous chloride and 20 cc. dilute hydrochloric 
acid to 1 liter) drop after drop until colorless, cool in water, add 
a 2^-per cent, solution of mercuric chloride in excess, stirring vig- 
orously (the precipitate formed should have a silky lustre), and 
titrate. 

b. Permanganate Method. — Decompose with hydrochloric acid 
as before, and add small pieces of zinc; when the ferric chloride is 
reduced, add cold water, transfer to beaker, dilute to 500 cc, add 
25 cc. cone, sulphuric aciil, and titrate. Another method: Decom- 
pose in a casserole with hydrochloric and sulphuric acids, heat 
until sulphuric acid fumes are given off ; cool, dilute slightly, re- 
duce with zinc, dilute, and proceed as before. 

3. Manganese? — Manganese is usually titrated with potassium 
l^ermanganate. Decompose the slag with hydrochloric acid in a cas- 
serole, add a few drops of nitric acid and 5 cc. of sulphuric acid, evap- 
orate till sulphuric acid fumes are given off ; cool, transfer to a 
beaker, dilute to 200 cc, boil, precipitate the iron with a cream-like 
emulsion of zinc oxide in large excess, filter into a flask, and wash. 

1 School of Mines Quarterly, vi., p. 163. 

- Engineering and Mining Journal, January 22, 1881. 

3 lies, School of Mines Quarterly, v., p. 223. 



SMELTING IN THE BLAST-FURNACE. 277 

Now either boil and titrate, shaking the flask, that the precipitate 
may settle and the finishing point be easily recognized, or filter 
into a ^-liter flask, fill up to the mark, take out 100 cc, transfer to 
a casserole, heat to a boil, and titrate. 

The Fe indicated by the cc. of the permanganate multiplied by 
0.2946 = Mil 

4. Lime. — This is titrated with potassium permanganate. De- 
compose the slag in a casserole as in determining silica, add to 
the hot silica filtrate a small excess of ammonia, dissolve the pre- 
cipitated iron, etc., in a hot solution of oxalic acid, using a slight 
excess; repeat the precipitation and solution, avoiding an excess of 
oxalic acid, boil for a few minutes, allow to stand 20 minutes in a 
warm place, filter, and wash with hot water till the filtrate does 
not taste of oxalic acid. Perforate the filter, wash the precipi- 
tate with hot water into a beaker, then dissolve with hot dilute 
sulphuric acid (some chemists use hydrochloric acid) any precip- 
itate that may adhere to the filter, and titrate hot with potassium 
permanganate. 

The Fe indicated by the cc. of the permanganate divided by 2 
= CaO. 

5. Baryta, Magnesia and Alumina are determined by the 
usual methods. 

6. Zinc. — This is determined by titrating with potassium ferro- 
cyanide (41.25 grs. to 1 liter), using uranium acetate (0.2 grs. 
in 100 cc. water) as an indicator. It is standardized with zinc 
oxide. 

a. Headdert 's Method. 1 — The ferrocyanide is standardized by 
dissolving pure zinc oxide in hydrochloric acid. Dissolve 1 gr. ore 
or slag in a casserole in hydrochloric acid, add nitric and sulphuric 
acids, evaporate till sulphuric acid fumes come off; cool, dilute 
precipitate once 2 with ammonia, filter, wash, add hyrochloric acid 
(the solution should not exceed 50 cc), precipitate the copper with 
aluminum foil, add from 8 to 10 cc. hydrochloric acid, dilute to 
225 cc, and titrate at 40 or 50° C. 

b. v. Shulz and Loitfs Method? — The ferrocyanide is standard- 
ized as follows. Dissolve 200 mgrs. recently ignited zinc oxide in 

1 Proceedings of the Colorado Scientific Society. 

- Coda, Oesterreichische Zeitschrift fur Berg- it. Hutten-Wesen, 1890, 
p. 123 : Engineering and Mining Journal, March 14, 1891. 
3 Proceedings of the Colorado Scientific Society. 



278 METALLURGY OF LEAD. 

a mixture of 5 grs. ammonium chloride, 15 cc. concentrated am- 
monia, and 25 cc. hot water; add 25 cc. concentrated hydrochloric 
acid, 1 gr. sodium sulphite and from 50 to *75 cc. hot water. The 
solvent for an ore (slag) is a saturated solution of potassium chlorate 
in nitric acid. It keeps well; the flask is best kept loosely covered. 
The mode of operating is: Dissolve 1 gr. of the substance in 25 cc. 
of the solvent in a 3^-inch casserole, keep this uncovered, warm 
gently till no more greenish vapors come off, cover with a watch- 
glass, and evaporate to dryness ; cool, add 5 grs. ammonium chlo- 
ride, 15 cc. concentrated ammonia, and 25 cc. hot water, boil one 
minute, break up any lumps with a stout glass rod, filter, and wash 
with hot water. If the filtrate be blue, precipitate the copper 
by adding 30 grs. of test-lead to the required 25 cc. hydrochloric 
acid and shake until it becomes colorless. Add 1 gr. sodium sul- 
phite and shake the solution (any sulphur that may be precipitated 
does not interfere with the result). In titrating, diluting the solu- 
tion to one liter and taking a measured quantity are avoided as fol- 
lows : Pour two-thirds of the solution into a beaker and titrate, 
passing the end point ; then add the rest of it, excepting a few cubic 
centimeters, and titrate again, using one cc. at a time until the 
end point is again passed ; finally rinse out the vessel holding the 
solution into the beaker, and finish the titration, testing after 
every 2 or 3 drops of the standardized solution. A determina- 
tion takes from 20 to 30 minutes. Gold, silver, lead, manganese, 
iron, sulphur, or the other constituents do not interfere with the 
results. 

§ 82. Wall- Accretions. — These accretions begin in lead fur 
naces just above the water-jackets and reach up to the feed-door. 
They sometimes come from the galena in the charge, or from 
galena formed artificially during the descent of the charge which 
adheres to and filters into the brick walls of tne furnace; but their 
principal origin is in the volatilization of lead, zinc, and their com- 
pounds, which takes place in the lower parts of the furnace. 
These fumes condense in the upper cooler parts. During their 
ascent they are partly oxidized, and the oxides may act chemically 
on the unaltered parts. Thus in wall-accretions may be found 
volatilized metal, metallic sulphides, arsenides, antimonides with 
their oxides, and secondary products. Any insoluble silicates found 
in the accretions come from parts of the furnace-lining or from fine 
particles of the charge. 



SMELTING IN THE BLAST-FURNACE. 



279 





Lead- 
ville. 1 


Lead- 
ville. 1 


Lead- 
ville.* 


Pueblo. 2 


Pueblo. 2 


Tarno- 
witz (sul- 
phide). 3 


Tarno- 
witz (ox- 
ide). 3 


Claus- 
thal.* 

1 


s 

Pb 

PbO 

Bi 


8.291 
47.491 
0.405 
trace 
0.0754 
trace 
trace 
6.977 
0.300 
trace 
6.754 
1.100 
2.887 
0.039 

trace 

6.965 

2.166 

trace 
5.361 
4.297 
1.672 

10.100 


2 600 
70 631 

6.297 
0.0007 

trace 

6*028 
3.456 

5*009 

2.*697 

5.593 

6.600 

2.830 
6.256 


2.725 
25.953 

6.0944 

53.*3'9*2 

1.381 

6*0*71 
6.056 

6.287 
6.035 

6.200 

0.159 

0.328 

1.577 

13.751 


25.98 
1.48 

6.0188 

none 
45.68 

6.71 

0.5*4* Mn. 
none 

trace 

6.20* 
8.13 


19.74 

37.48 

6*0*824 

none 
38.99 

6.94 

none 
none 

6*12 

6.30 
6.57 


13-9 
20-32 

6.005 

30-27 
C. 2. 


'24* 
6.665 

*60* 


14.67 

82.71 

trace 
2.03 

1 .79 


Ag 

Au 

Cu 

Zn 

ZnO 

Cd 


Fe 

Fe 3 

Mn'oOj 

As 


\ 


9-6 

18 


Aso0 3 

Sb". 


SboOa 

Sn"! 

p 2 o, 

CI, Br 

C0 2 

CaO 

MgO 

A1 2 3 

SiO s 

O 









1 Emmons, "Geology and Mining Industry of Leadville," p. 727. 

2 Dewey, Bulletin No. 42, "United States National Museum," p. 54. 

3 Dobers and Dziegiecki, Z.f. B.- H.-, u. S.- W. i. P., xxxii., p. 102. 
* Mezger, Berg- und Hiittenm. Ztg., 1853, p. 52. 



The Clausthal analysis, where galena is smelted raw, represents 
a crystallized wall-accretion, consisting principally of galena. At 
Tarnowitz, where slag-roasted galena rich in blende is smelted with 
the gray from the reverberatory furnaces (§ 44), two kinds of ac- 
cretions form, one a sulphide principally black, the other an oxi- 
dized compound having a greenish color. In the three analyses by 
Guyard, of Leadville, accretions where carbonate ores formed, if not 
the whole, at least the major part of the charge, shows a great A r a- 
riety in composition. The lead, for instance, is present as metal, 
as sulphide, and as oxide ; zinc, arsenic, and antimony as sulphides 
and as oxides. 

The method of removing these accretions while the furnace is 
running and their treatment have already been discussed (§ 66). 



280 



METALLURGY OF LEAD. 



§ 83. Hearth-Accretions or Sows.— In a furnace with 
an Arents' syphon tap these unwelcome products form on top of 
the- lead below the tuyeres; in furnaces where lead, speise, and 
matte are tapped from the bottom, they form there. They result 
from a faulty charge or from a lack of fuel, and are mixtures of 
slag, speise, matte, metallic iron, metallic lead, coke, and charcoal. 
The metallic iron results from the reduction of ferric oxide, some 
of it being held in solution by melted matte and dropped when 
this cools. The iron of a sow is generally carbonized and contains 
silicon and phosphorus. 



Fe 

Pb. 

Ag. 

Au. 

Cu. 

As. 

Sb. 

Mo. 

Ni. 

Co.. 

Zn.. 

Mq 

S... 



p 

Graphite 

Combined carbon. 

Si, slag loss 

CaO 

SiOo 



Leadville. 1 


Pribram. 2 


Pribram. 2 


72.828 


48.685 


38.875 


18.793 


7.309 


20.140 


0.1149 


0.060 


0.160 


0.00003 






trace 


0.199 


1.107 


5.083 


8.446 


2.389 


trace 


0.450 


0.735 


0.161 




.... 


0.045 


trace 


trace 


trace 


trace 


trace 


trace 


3.610 


5.417 


0.015 






0.650 


13.760 


18.466 


0.109 




.... 


0.750 


.... 


.... 


0.550 






0.900 








0.166 


0.780 




15.850 


9.300 



1 Emmons, Op. cit., p. 723. 

2 Balling, Berg- u. Hiittenm. Ztg. 



r, P . 4i9. 



It does not usually pay to work up a hearth-accretion; it is 
thrown over the dump or buried, being an eye-sore. Flechner 1 
suggests several methods of working furnace-sows. The following, 
used at the nickel works of Schwerte in Westphalia (Prussia), is 
of interest. Furnace-sows containing from 75 to 85 per cent, iron, 
5 to 8 copper, 3 to 6 molybdenum, 2 to 4 nickel-cobalt, and weigh- 
ing from 500 to 600 pounds apiece, are gradually melted down 
with coke on the bottom of a blast-furnace, the melted parts run- 
ning out continuously. In this way a large crust is easily reduced 
in size, and can then be added again to the ore-charge, where it will 
be taken up by the speise and the matte. Any mechanical means 



1 Oesterreichische Zeitschrift filr Berg- u. Hiitten-Wesen, 1889, p. 196. 



SMELTING IN THE BLAST-FURNACE. 



281 



of breaking up a hearth-accretion is sure to cost more than will be 
recovered from re-smelting afterward. 

§ 84. Furnace Cleanings and Furnace Refuse.— Furnace 
cleanings and refuse are a mixture of fire-brick, metal-bearing 
compounds, fuel, etc., obtained in cleaning out a blast-furnace 
when blown down. They are assorted; the waste goes to the slag- 
heap, the valuable part is added to the ore-bed. 

§ 85. Flue-Dust or Chamber-Dust.— This product is as 
important as it is unwelcome. It consists of fine particles of the 





Wyandotte, Mich" 


Pueblo. 


Hartz 
Mts. 


Ems. 


Freiberg. 


Sheffield. 


Pb 








34.8 
18.0 

2.9 

1.0 
1.5 


60.48 


67.04 


35.2 


27.90 




PbS2.25 
68.35 


PbO 

Pb.,Si0 4 
Zn~. ... 


19.91 


23.77 


37.65 


44.80 














3 17 


4.22 


5 28 


49.50 






ZnO .... 
Cu 


0.09 


trace 


5.32 
trace 


4 80 
H-Bil.52* 


1.80 


trace 


trace 


trace 
1 30 

28.3 




CdO 






As s 








J- 3.0 


0.24 
0.42 
0.003 


0.16 
0.31 
0.003 


1.60 


t 3.03* 





Sb. . 








Ag 

Au 


0.286 


292 


0.04 














Fe 








1 
4.5 














Fe 2 O s . . 
Al 2 O s . . 
NiO ... . 
CoO .... 
CaO .... 
MgO.... 
Si0 2 .... 
s 


1 14.43 

0.08 
0.09 
8.74 
366 
16.11 


1 18.54 

trace 
trace 
6.62 

' 13.06 


24.98 
1.31 


(2.12 


[l.OO 


1.57 




trace 
10.00 


'"546" 






















5.26 
none 
8.63 
2.53 
1.61 

ill. 20 

2. 




1.15 


0.61 


1.01 
25 
6.19 


7.00 




2.63 




12.3 

7.8 
2.8 


' 6.22 

14.78 


' 5.42 
14.07 




9.00 


2.25 
16.84 


S0 3 

HoO.... 
C6 2 . . . . 
C 

Reference : 


9.30 
i 1.28 
19.23 

l. 


8.85 

[3.76 

22.14 

1. 


3.38 


13.00 


28.81 


2.5 

3. 


8.00 
4. 


5.80 

4. 


1.17 
5. 








5. 


6. 


6. 



* As oxides . 

(1) Curtis, Trans. A. I. M. E.. ii., p. 95; (2) Dewey, Bulletin No. 42, "United States National Mu- 
seum," p. 53 ; (3) Balling, Metallhiittenkunde, p. 87 ; (4) Freudenberg, Op cit., p. 19 ; (5) Hering-, Op. cit. t 
p. 34 ; (6) French, Engineering and Mining Journal, January 17, 1880. 

charge carried out of the furnace by the ascending gases, and of 
metals and their compounds that have been volatilized in the lower 
parts of the furnace and not been again condensed in it, but settled 
out only while passing through condensing flues or chambers. 

The flue-dust from blast-furnaces has generally a dark color, 
which is caused by the admixture of finely divided fuel. It is 
black when charcoal forms even a small part of the fuel. The 



282 METALLURGY OF LEAD. 

amount of flue-dust formed depends upon a variety of causes. 
Fine ores or fluxes are carried away easily by the current of gases; 
charcoal, being friable, makes dust; soft coke is broken up to some 
extent, and causes mechanical losses; and the manipulation of the 
furnace, affecting the descent of the charge, has a very great influ- 
ence. Thus careful feeding and cutting out of wall-accretions will 
reduce the forming of flue-dust to a great extent. Then, a high 
temperature in the smelting zone causes much volatilization; in the 
same way a high blast will cause much vapor to be carried out of 
the furnace if the quick ascent of the gases be not checked by the 
form of the furnace (by having boshes). Extreme figures of the 
amount of flue-dust formed are 0.8 and 15 per cent, of the weight 
of the ore charged. 1 An average figure is 5 per cent. 

The different methods of condensing flue-dust 2 may be classed 
as wet and dry. With the former the gases are either drawn 
through water alone, 3 or they are pressed through one or more 
horizontal filters, 4 through which fine sprays of water trickle, car- 
rying with them the condensed vapors and filtered solid particles. 
Water condensation, w r hile still found in England, 5 is not much 
used elsewhere, as it is on the whole imperfect, and the apparatus 
and their maintenance costly. Dry condensation is the almost uni- 
versal method used. It consists in cooling the gases and retarding 
the velocity of the current. Dry filtering is found in a few in- 
stances and electrical condensation has been experimented w T ith. 

The cooling of gases is essential to an effective condensation, 
as the single particles, being brought closer together, unite more 
easily into flaky masses, which then settle out. While the gases 
from a lead blast-furnace require very little cooling, the vapors and 
dust coming from reverberatory furnaces used in smelting (§ 39) 
and slag-roasting (§ 56), and in desilverizing base bullion (§97 and 
105), are apt to be lost, if their temperature is not reduced. 

An arrangement Avhereby the gases are cooled by air alone is 
shown in the following sketch, 6 Figure 164. It represents a hori- 
zontal sheet-iron flue, with small sliding doors at intervals of two 

1 Hahn, " Mineral Resources of the United States," 1882, p. 344. 

2 Hering, C. A., "Die Verdichtung des Huttenrauches," Stuttgart, 
1888. 

3 Eilers, Trans. A. I. M. E., iii., p. 310. 

4 Percy, "Metallurgy of Lead," p. 442. 

D Rosing, Z. f. B., H. u. S. W. i. Preussen, xxxvi., p. 103. 
6 Eilers, Trans. A. I. M. E., iii., p. 309. 



SMELTING IN THE BLAST-FURNACE. 283 

feet, suspended by means of iron rods from wooden trestles. It 
was used at the Richmond Works, Nevada, was 800 feet long, and 
ended in a wooden stack 40 feet high. 

A sheet-iron flue of a different form but filling the same purpose, 
used at the Ems ■ Smelting "Works (Prussia), is shown in Figures 
165 and 166. The flues proper, 1 by 0.75 meters (3 feet 3| inches by 
2 feet 5| inches) and 1 meter square (3 feet 3| inches), made of 



FIG. 164. — SUSPENDED SHEET IRON FLUE. 

T<r-inch iron, have triangular projections, the base of the triangle 
being 6-J- feet long and the height 3 feet 3 f inches. 

To make the cooling more effective and thus shorten the flue, 
water has been used in different ways. In 1877 Hagen 2 introduced 
in Freiberg his water-cooled flues made of sheet-lead (Figures 167 

1 Egleston, Ibid., xi., p. 410, plate ii. 

2 Freiberger Jahrbuch, 1879, p. 151 ; Hering, Op. cit., p. 20, and Pri- 
vate notes, 1890. 



284 METALLURGY OF LEAD. 

and 168). The roof (5 feet 3 inches) and sides (8 feet 2f inches) are 
suspended from a wooden framework in the same way as with a 
sulphuric-acid chamber. The top c is cooled by a slight flow of 
water d. This circulates along the sides through elliptical pipes 
b, soldered to the flue and to each other. Figure 16V, a vertical 
longitudinal section through the pipes £>, shows the circulation of 
the water. This arrangement permits cooling with less water 
than would be required if it overflowed from the roof down the 
sides. 

Another arrangement has lately been introduced at Freiberg 
by Richter. 1 It consists in letting the gases ascend in a set 
(live) of wooden towers, lined with -J-inch sheet-lead. These are 18^- 
feet high and 6^- by 10 feet in the clear. Water trickles down in the 
towers, each of which contains ten rows of small roofs, supported by 
lead-coated rails. Each roof consists of three sheets of hard lead, 
3 feet 3-J- inches long and F 5 ¥ inches thick, bent to the form of a V, 





FIGS. 165-166.— SHEET-IRON FLUES WITH TRIANGULAR PROJECTIONS. 

and having incisions 1-| inches deep at the lower border, in order 
that the water may run off freely and leave the necessary inter- 
stices for the ascending gases. The roofs are placed so that the 
discharge of one set strikes on the ridge-poles of the next one below. 
The water for the five towers is drawn from two large tanks, and a 
short distance above the lowest row of roofs is a coarse grating, 
w T hich is to distribute the water evenly in order that the gases, on 
striking the first roof, may be cooled as much as possible. The 
temperature of the gases on entering the tow r ers is from 100 to 115° 
C, and on leaving from 40 to 60° C. ; they then pass through a 
leaden flue 1,6*73 feet in length before entering the main stack, 
w^hich is 453 feet high. Each tower can be separately shut out 
from the main flue, which is done at certain intervals, to remove, 

1 Freiberg er Jahrbuch, 1889, p. 57; Berg- u. Huttenm. Zeitung, 1890, 
p. 129 ; Engineering and Mining Journal, Feb. 15, 1890, and Private notes, 
1890. 



SMELTING IN THE BLAST-FURNACE. 



285 



by means of a stream of water, the flue-dust that collects beneath 
the reofs. 

The main effect of this apparatus is due not so much to the 
larger surface it offers to the gas-current as to the fact that it cools 
the gases. Five times as much flue-dust is collected with it as be- 
fore. 

Other arrangements for cooling gases are those by Schlosser and 
Ernst, 1 who let cooled solutions circulate in coils of iron pipe 
placed at certain intervals in the main gas-flue. Another plan 2 is 
to make gases pass through a circular tower in which iron pipes, 



Fig. 167 



d& 




b b 

WATER-COOLED FLUE OF SHEET-LEAD. 



with water circulating in them, are suspended. Any dust adher- 
ing to the pipes is removed periodically by a jet of water, intro- 
duced through a movable central pipe having nipples at certain in- 
tervals. 

In settling out flue-dust the most important thing after cooling 
is the retarding of the velocity of the air-current and the exposing 
of a large condensation surface to it. It has been found that the 
point of least friction is on the centre line of a flue at to of its 

1 Berg- u. Huttenm. Ztg., 1885, p. 464; 1887, p. 134. 

2 Wagner's Jahresberichte, 1889, p. 300. 



236 METALLURGY OF LEAD. 

height. Here the air has its greatest velocity. This diminishes 
very little upward, but so much downward that there may even be 
a small counter-current at the bottom of a flue. The speed of the 
air-current can be retarded by increasing the volume of the flue, 
by changing its direction or by increasing its surface. 

Increasing the size of flues through which the gases pass has 
always been considered a very effective means of settling out flue- 
dust. On entering an enlarged chamber the velocity of the current 
is gradually slackened to the point where the draught near the exit 
begins to show its effect ; then there is a gradual increase of ve- 
locity. On this account a comparatively small part of the enlarged 
flue or chamber is really useful in the settling out of the dust. 

The reason that a change in the direction of the current is also 
not so effective as might be expected, is that, though at the turning- 
point the speed of the current may be slackened for a short dis- 
tance, the pressure of the air behind and the draught in front will 
quickly restore it to the normal rate. 

It is Freudenberg's * great merit to have discovered that an in- 
crease of surface is the most effective means of settling particles 
held in suspension by a current of air, and that the amount of 
flue-dust settled out stands in direct proportion to the area of sur- 
face with w T hich it comes in contact. This increase of surface 
means increase of friction between the stationary surface and the 
moving Current of gas, a consequent retarding of its velocity, and 
with it a settling out of dust particles. The surface also attracts 
these, if it is cold. 

Aitken 2 has shown that a hot surface repels them, especially 
if moist. This explains how the introduction of vaporized or finely 
divided water 3 into the air- current has not proved so effective as 
was anticipated. 

In order to increase the surface, Freudenberg suspended thin 
sheet-iron plates parallel with the air-current; and to prevent the 

1 Freudenberg, " Die auf der Bleihiitte bei Ems zur Gewinnung des 
Flugstaubes getroffenen Einrichtungen," Ems, 1882. Abstract, Engineer- 
ing and Mining Journal, July 1, 1882. Egleston, Trans. A. I. M. E., xi., 
p. 379. Stetefeldt, "Comment on Freudenberg's Plates," Engineering 
and Mining Journal, July 28, 1883. 

2 Proceedings of the Royal Society of Edinburgh, xxxii., p. 239; 
Wagner s Jafiresberichte, 1884, p. 1307. 

3 lies and Keiper, Engineering and Milling Journal, February 27, 



SMELTING IN THE BLAST-FURNACE. 



287 



dust from being carried off, after it had once settled out, he placed 
partitions across the bottom, reaching nearly to the hanging plates. 
In Figures 169 and 170, representing a vertical and a horizontal 
section of a flue, are seen the sheet-iron plates B, 3-j- inches apart, 
having pieces of iron D, bent to the form of a hook, riveted to 
their ends. By these they are suspended from pins passing through 
rectangular cross-bars L, which, twisted flat at both ends, reach 
1-J- inches into the side-walls A of the flue. The flue-dust collect- 
ing on the plates falls off when it has grown to a certain thickness. 
Every 16 or 20 feet cross-partitions E, 7j inches high, protect the 



Fig. 169. 



Fig. 170. 




Scale i: 80 






en C| 




FLUE WITH FREUDENBERG PLATES. 



lower part of the flue from the air-current. They are bolted to 
cast-iron supports C. 

The investigations of Freudenberg form the basis for the 
construction of most modern dust-chambers. Rosing 1 has sub- 
stituted iron wires for Freudenberg plates. By a shaking arrange- 
ment the dust can be more easily removed from the wire than from 
the plates. In Hering's book, already quoted, are given numerous 
sketches of condensation chambers that have been projected or 
built since Freudenberg's results became generally known. 

Filtering is not much used, owing probably to the fact that 



1 Patent No. 432,440, July 15, 1890. 



288 



METALLURGY OF LEAD. 








., 



i 



nearly all the dust can be recovered with- 
out it. The effectiveness of filtering is 
certain. Whether it pays to draw off all 
the gases from a number of blast-furnaces 
with fans will be known when the results 
obtained by lies at the Globe Smelting 
and Refining Co.'s Works, at Denver, be- 
come public. The Lewis-Bartlett filtering 
process has already been described (§ 51). 

The use of electricity 1 in settling flue- 
dust has not been so successful as the first 
experiments seemed to promise. In closed 
chambers static electricity quickly clears 
dust-laden air ; when, however, the air 
is in motion, electricity produces no effect. 

As regards the arrangement of the 
plant, it is important to have a dust-cham- 
ber near the furnace, as most of the heavy 
dust is settled out quickly. The chamber 
will lead into a flue (or flues) with Freu- 
denberg plates, where the gases will drop 
most of their dust, and then pass off 
through a chimney. Sometimes it will 
be necessary to have a fireplace at the 

1 Hutchings, Berg-u. Huttenm.Ztg., 1885, 
p. 253 ; Engineering and Mining Journal, 
May 8, 1886. Rosing, Berg- u. Huttenm. Ztg., 
1885, p. 290. Bartlett, Engineering and Min- 
ing Journal, March 13, 1886. 



SMELTING IN THE BLAST-FURNACE. 



289 



foot of the chimney to create the required draught. Sometimes 

a fan in the flue may be needed to suck and push onward the gases. 

The dust-chamber of the Montana Smelting Co.'s Works, at 

Great Falls, Mont., shown in Figures 171—177, may serve as an ex- 




Kig. 174= 



<^Z^> Fig.lTS 



Fig. 177 





FLUE AND DUST-CHAMBER OF THE MONTANA SMELTING CO.'S WORKS. 



ample of the manner in which the dust is now collected at most 
lead smelters. A general plan and elevation of the works has been 
given in Figures 82 and 83. In Figure 171 is shown a horizontal 
section of the main flue a, which runs along the back of the furnace 
floor. It receives the dust- laden gases from the blast-furnaces 



290 



METALLURGY OF LEAD. 



through openings (j in the roof. The gases pass from the flue to 
the chambers e, and then into the stack. 

If Ffeudenberg plates were used, they would be suspended in 
that part of the flue reaching from the chamber e to the first 
"blast-furnace g. 

The main flue has at intervals of 9 feet 4 inches small openings 
b (Figures 171, 172, 173), placed diagonally opposite each other, 
through which the flue-dust is raked out periodically into the 
shallow pits f. They are closed with hinged iron doors. Every 
5Q feet (the distance between the centres of two blast-furnaces) 
there is a manhole c (Figures 171 and 174), which is bricked up 
with a half course of bricks. These are easily removed when 
the flue is to be entered. For each blast-furnace will be found 




I BLAST FURNACE STACK 

I MONTANA SMELTING CO. 



BLAST-FURNACE STACK OF THE MONTANA SMELTING CO.'S WORKS. 



in the side wall a small opening d (Figures 171, 175-177). This 
is connected by an underground flue with the stove that heats the 
cast-iron lead-pot placed next to the lead-well (§ 64). From the 
flue a the gases pass four dust-chambers e, having vaulted arches 
running parallel to that of flue a. The area of the passage from 
the main flue to the first chamber and of the one between the 
single chambers is somewhat contracted. The gases must thus 
pass the openings in the partition walls with an increased velocity, 
which diminishes again as soon as they enter the next chamber. 
The last chamber ends in the stack, of which Figures 178 and 
179 give the details. 

Dust-flues and chambers are usually built of brick, and require 
strong walls and close binding to keep them air-tight. They are 



SMELTING IN THE BLAST-FURNACE. 291 

therefore expensive, and are often made too small and too short to 
serve their purpose satisfactorily. A lightly constructed and rela- 
tively cheap dust-flue, 6 feet 6f inches by 9 feet 10 J inches and 
1,640 feet long, was erected two years ago at the Yictor-Friedrich 
Smelting Works 1 in the Hartz Mountains, according to the Monier 2 
system. It has so far given complete satisfaction. The principle 
of the Monier system is to reinforce the defective tensile strength 
of cement concrete by incorporating in it a coarse network of iron 
wire. It is still in the process of development, but would seem 
to be well adapted for dust-flues and dust-chambers, as the thin- 
ner walls must materially assist the cooling of the gases and thus 
promote the condensing of vapors and the settling out of dust par- 
ticles. 

The success of the new flues and chambers just erected at the 
works of the Grant and Omaha Smelting and Refining Company, 
Denver, Col., 3 will be of great interest. They are built under- 
ground, the walls being made of common hollow brick opening at 
one end into the air and at the other into a stack, which draws 
fresh air continually through them and thus keeps them cool. 

§ 86. Treatment of Flue-Dust.* — The question how to treat 
flue-dust with a minimum of loss is a difficult one. Many sugges- 
tions have been made and various methods tried. 

To wet down flue-dust and put it back into the furnace is of no 
use, as when dry it will simply be blown out again. 

An improvement on this is to mix the flue-dust with from 8 to 
10 per cent, of slacked lime and form it into bricks. These, when 
hardened, are added to the furnace charge. Cahen 5 gives an exam- 
ple of mixing flue-dust (with 44 per cent, lead) and slacked lime. 
The bricks formed were smelted alone in the blast-furnace with 74 
per cent, of slag, and only 9.60 per cent, of the lead charged was 
lost. If the bricks are to stand any rough handling, or the weight 
of the charge in the blast-furnace, they must not be made in a pug- 
mill, but in a machine and under considerable pressure. They 
must then be dried slowly with artificial heat. Harbordt, 6 having 

1 Berg- und Hiittenmannische Zeitung, 1891, pp. 175, 439 ; 1892, p. 66. 

2 Engineering and Mining Journal, June 6, 1891 ; Berg- und Hiitten- 
mannische Zeitung, 1891, p. 175. 

3 Iron Age, May 19, 1892. 

4 Ites, Eng. and Min. Journal, January 30, February 6, 27, 1886. 

5 " Metallurg-ie du plomb," Liege, 1863, p. 102. 

6 Private communication, July, 1891. 



292 METALLURGY OF LEAD. 

found bricking with lime unsatisfactory, substituted clay. The 
bricks, simply air-dried, can stand considerable handling, and do 
not crumble in the blast-furnace. Many different ways of bricking 
tine ore and flue-dust have been tried. Hahn 1 found that flue-dust 
mixed with a solution of ferrous sulphate and formed into bricks 
(by hand) became so hard when sun-dried that it would stand 
much handling. Perhaps an addition of slacked lime might prove 
effective, as it precipitates the ferrous oxide, which, on being ex- 
posed to the air, becomes oxidized and forms hard lumps. 

The writer observed, while roasting blende-bearing galena to 
obtain zinc sulphate for leaching, that, after removing as much of 
the zinc as was possible, the residual ore, when dry, became so hard 
that it required strong blows with a sledge to break it. 

Church 2 made brick of fine ore and flue-dust in Tombstone, 
Arizona, by using pan-slimes, settled in tanks, as binding material. 
The bricks were made without pressure, and sun-dried. These 
slimes contained 85 per cent, quartz and from 2 to 3 percent, clay, 
the remainder being calcite, manganese, iron oxides, sulphides, 
and lead carbonate. Church attributes the binding quality of the 
slimes entirely to their fineness. It is a well-known fact that ma- 
terial when ground very fine will often show adhesiveness it did 
not possess before. 

Peters 3 used at the Parrott Works, Butte, Montana, a brick-ma- 
chine that exerted a pressure of four tons to the square inch, and 
was capable of making forty bricks a minute, each Aveighing five 
pounds when dry. The ore to be compressed was roasted pyrite, 
and no binding material was used. The bricks had to be dried 
slowly by artificial heat to be suitable for the blast-furnace. He 
states that with an addition of from 2 to 4 per cent, lime or clay, 
slow drying could be dispensed with. 

It is to be noted that in all bricking experiments the results 
obtained with fine ore cannot be simply adopted for flue-dust, as 
this contains carbonaceous matter which counteracts some of the 
binding property of adhesive substances. The bricking of flue- 
dust, even if imperfectly done, has always the advantage that the 
bricks will be carried some distance down with the descending 
charge before they crumble. The major part will thus be held back 

1 "Mineral Resources of the United States," 1882, p. 344. 

2 Trans. A. I. M. E., xv., p. 611 ; Engineering and Mining Journal, 
August 22, 1885. 

3 " Modern Copper Smelting," 1891, p. 279. 



SMEL TING IN THE BLAST-FURNA CE. 293 

by the charge above, and prevented from passing off again into the 
dust-chamber. The disadvantage of bricking is that, oxidized lead 
and carbonaceous matter being in intimate contact, finely divided 
metallic lead is formed in the upper part of the furnace and causes 
loss by volatilization. 

For this reason, as well as on account of the cost of making 
bricks, it is more common to melt the flue-dust in a special rever- 
beratory furnace (see ante) or in the fuse-box of the roasting fur- 
nace, where it is added in quantities of, say, 100 pounds to the 
roasted ore. It is charged before the ore is drawn from the roast- 
ing-hearth, and is thus covered by the roasted ore, so that little 
metal is carried off mechanically. It cannot, however, be denied 
that slagging flue-dust causes some loss in lead by volatilization; 
the loss in silver is inconsiderable. 

§ 87. Analytical Determinations. — The assay of flue-dust 
requires no special discussion, as it contains nothing which cannot 
be found under the heads of speise (§ 74), matte (§ 79), and slag 

(§ 81). 

§ 88. Losses in Smelting. — The losses in smelting are due 
to slagging or to particles of metal being carried off in by-products 
and not recovered. As has already been stated (§ 53), ore slags 
should not contain over f per cent, lead and -J-ounce silver to the 
ton with 300-ounce bullion; but they often contain over 1 per 
cent, lead and about 1 ounce silver, and are considered satisfac- 
tory. Special causes, such as the presence of foreign substances 
having a deleterious effect (§ 54), may make slags run still higher. 
The quantity of by-products (speise, matte, flue-dust) formed has 
an important influence on the output of lead and silver, as they have 
to be roasted and resmelted several times, and each of these opera- 
tions causes an unavoidable loss in metal. It is difficult, therefore, 
to give an average figure of the lead and silver recovered in smelt- 
ing in the blast-furnace. It may be said that a total loss of 7 per 
cent, of lead represents very good work, the lead assays being made 
in the dry way. Losses of twice the amount are, however, not un- 
common. With silver the output is generally over 95 per cent., 
the silver being also determined in the dry way. If a smelter can 
afford to pay (§ 35) 95 per cent, of the assay value of the silver, 5 
per cent, or less must be the maximum he endures with the ordi- 
nary run of ores. 

§ 89. Cost Of Smelting. — The cost of smelting a self -fluxing 
ore at the large smelters of Denver and Pueblo is $4.50 per ton, 



294 METALLURGY OF LEAD. 

as previously stated (§ 35). It may go as high as $5.00, but not 
higher. The figure includes all incidental expenses; in fact, is the 
total cost. 

The following estimate is by C. Henrich, who has often been 
quoted in the preceding pages. It is for a single furnace, 33 by 
84 inches, based on Colorado data, and assumes that the furnace 
puts through 48 tons of ore in 24 hours and requires 16 tons of 
flux. 

MATERIALS : 

16 tons flux @ $3.00 $48.00 

11 T 7 T tons coke (18 %) @ $12.00 , 132.77 

3 cords wood @ $6.00 18.00 

Supplies, etc 12.50 

$211.27 

PAY-ROLL : 

Superintendent -. $11.70 

Assayer. ... 5.00 

Foreman 5.00 

Weighmaster 2.50 

2 engine-men @ $3.00 6.00 

3 furnace-men @ $3.00 9.00 

4 slag-men @ $2.50 10.00 

3feeders@ $3.00 9.00 

4 charge-wheelers @ $2.50 10.00 

8 laborers @ $2.00 16.00 

2 inside laborers @ $2.25 4.50 

1 sampler 2.50 

2 bullion-men @ $2.00 4.00 



95 20 
$306.47 
Delays in repairing furnace, 5 per cent. 



For unforeseen expenses, 5 per cent. . . ) 

;. ) 



Cost of smelting 48 tons of ore $337.12 

Cost of smelting one ton of ore 7.03 

It will be seen that he assumes 8-hour shifts for furnace-men 
and feeders, 10-hour shifts for day laborers, and 12-hour shifts for 
weighmaster and engineers; 10 per cent, is added to the cost for 
unforeseen expenses and repairs of the furnace. This is as accurate 
a general statement as can be made. 



PAET III. 

DESILVERIZATIOfl OF BASE BULLION, 



OHAPTEE IX. 

PATTINSON'S PROCESS. 

§ 91. Introductory Remarks. — The process is based on the 
fact discovered by Pattinson in 1833, that, if silver-bearing lead is 
melted and cooled down almost to its fusing-point, crystals of lead 
will separate which are much poorer in silver than the original lead. 
If they are removed and the process is repeated, always adding 
fresh lead of the same tenor in silver, a large quantity of market 
lead low in silver will result, and a small amount of enriched lead 
ready to be cupelled. By the repeated meltings and crystallizations 
many of the impurities will also have been collected in drosses, and 
the market lead become purified. 

The fact that lead low in silver solidifies before the enriched 
lead still lacks a satisfactory explanation. Pattinson ' observed 
that on heating carefully a bar of lead that ran low in silver, until 
a few drops of metal oozed out, these were richer in silver than the 
residual lead, while with a bar running high in silver just the 
reverse was the case. Reich 2 measured the melting-points of 
different lead-silver alloys, and found that lead with 1.89 ounces 
silver per ton melts at 321° C; with 139.03 ounces, at 309°; and that 
with 656.23 ounces silver the lowest melting-point was reached. If, 
however, the lead contains much silver, for instance 33 or 50 per cent., 
its melting-point is much higher than that of lead free from silver. 
The inferences to be drawn from this are that Pattinson's process 
is adapted only for low-grade bullion, and that the enriching of the 
liquid lead can be carried only to a certain degree. The following 
table, by Reich, 3 shows how far the silver can be concentrated in 
the lead : 

1 Percy, "Metallurgy of Lead," London, 1870, p. 137. 

2 Berg- und Huttenm. Ztg., 1862, p. 251. 

3 Ibid. 



DESILVERIZATION OF BASE BULLION. 



297 



at the Ems Smelting and Refining Works, Prussia, where the three 
processes were used one after another. To these have been added 
the amount of lead to be cupelled and the traces of gold. 





Cupellation 
Process. 


Pattinson's 
Process. 


Parkes' 
Process. 


Cost 


3 

100 

6 

0.2 
0.73 
lost 


U 
13 
2 

0.05 
0.58 
lost 


1 
5 
1 

0.015 

0.17 

recovered 


A.niount of lead to be cupelled . . 


Loss in lead and silver 


Impurities remaining in the lead, per cent. 
Silver remaining in the lead, ounces per ton 
Traces of ^old 





The task of modern desilverizing works (or refining works, as 
they are also called), is not only to separate effectively and cheaply 
the precious metals from lead, but also to make out of a base bullion 
containing from 95 to 98 per cent, of lead, a refined lead of not less 
than 99.9 per cent, of lead and salable products of the impurities 
contained in it, such as copper, tin, arsenic, and antimony. 

In the following chapters, cupellation, being now only an auxil- 
iary to Pattinson's and Parkes' processes in desilverizing base 
bullion, although still an independent process in the manufacture 
of litharge, will be treated last. The three processes will be dis- 
cussed in the following order : 

Pattinson's Process. 
Parkes' Process. 
Cupellation. 



PART III. 
DESILVERIZATION OF BASE BULLION. 

§ 90. Introductory. — The final separation of silver and lead is 
universally accomplished by the process of cupellation. Keith 1 a 
desilverized base bullion on a working scale by means of electricity, 
but did not make a pecuniary success of it. Roesing 3 experimented 
at Tarnowitz, Silesia, in oxidizing lead in a Bessemer converter 
lined with basic refractory material. He worked with charges 
weighing 13,200 pounds, and enriched lead from 12.4 to 1863. 
ounces silver per ton, the fumes collected assaying 75 per cent, lead 
and 2.5 ounces silver per ton. He also refined, in a few minutes, 
desilverized zinc-bearing lead. In what way this new idea can be 
applied to desilverization so as to compete with the present methods 
remains to be seen. Formerly all argentiferous lead was cupelled, 
but this was found to have disadvantages, prominent among which 
are the cost and the loss in metal, the limit being very soon 
reached, where the separation of silver from lead ceases to pay. 
This is with base bullion assaying about 30 ounces silver to the ton. 
Below this the silver recovered will hardly pay for the labor, fuel, 
and material required, the loss in metal, and the impurity of the 
lead obtained from the reduction of the litharge. It becomes, 
therefore, necessary to concentrate the silver in a smaller amount 
of lead before cupelling. The processes of Pattinson and Parkes 
do this successfully. 

The progress made in desilverizing during the last sixty years 
is well illustrated by the following table. The figures published 
by Hermann 4 are derived from the actual working-results obtained 

1 Engineering and Milling Journal, July 13, 27, 1878 ; June 3, 1882 ; De- 
cember 15, 1883, p. 372. Trans. A. I. M. E., x., p. 312 ; xiii., p. 310. 

2 Hampe's criticism : Zeitschrift fur Berg-, Hutten-, u. Salinen- Wesen im 
Preussen, xxx., p. 91. Engineering and Mining Journal, March 18, 1882. 

3 Revue universelle des mines, 1892, xvii., p. 110 ; Iron, March 25, 1892 ; 
Berg- und Huttenm. Ztg., 1892, p. 102 ; Engineering and Mining Journal, 
April 16, 1892 ; Stahl und Eisen, 1892, p. 370. 

4 Berg- und Huttenm. Ztg., 1883, p. 382. 



PATTINSON'S PROCESS. 



299 



Ounces Silver per Ton. 


In the molten lead be- 
fore crystallization. 


In the crystals. 


In the liquid, lead. ■" 


205.33 
213.49 
281.34 

288.16 
420.57 
609.57 
615.15 
643.40 


113.74-135.91 
92.75-109.08 
119.58-198.33 
113.74-181.99 
198.91 
586.53 

503.99-646.31 
645.15 


298.95 
313.83 
422.91 
446.24 
560.57 
659.15 
655.65 
660.32 



The process has to stop when the liquid lead assays from 600 to 
650 ounces silver per ton. In practice the concentration is stopped 
when the liquid lead assays from 450 to 500 ounces, as the nearer 
the silver-contents approach the 650 ounces silver per ton, the 
smaller become the crystals, and the more difficult is it to drain 
off the liquid lead, especially as this also tends to solidify at the 
same time. 

The process of concentrating the silver in a small amount of 
lead may be conducted according to two systems, called the method 
by thirds and the method by eighths. In the first of these systems 
§ of the lead contained in the kettle is withdrawn in the form of 
crystals, while -J remains behind as liquid lead. The crystals will 
then be half as rich, and the liquid lead twice as rich, as the original 
bullion. In the second system the bullion in the kettle is divided 
into -J crystals and J- liquid lead, and the silver-contents of the 
crystals is \ as much, and of the liquid lead 3 times as much, as that 
of the original bullion. The latter method is, therefore, to be 
applied to low-grade bullion. Stetefeldt 1 tried to find a general 
mathematical formula which would show the proportions in which 
leads of different contents in silver should be divided to attain, 
with as few crystallizations as possible, a market lead of a certain 
tenor in silver and an enriched lead. In practice the two methods 
by thirds and by eighths have become standards, especially the 
former. A variation of the method by thirds, the one with inter- 
mediary crystals, 2 aims to reduce the number of crystallizations. 
The contents of the kettle are divided into | crystals and -| liquid 
lead ; the liquid lead, however, undergoes directly a second crystal- 

1 Berg- und Euttenm. Ztg., 1863, pp. 64, 69, 77. 

2 Stetefeldt, Berg- und Euttenm. Ztg. 1863, pp. 297, 381. 



300 DESILVERIZATION OF BASE BULLION. 

lization, so that intermediary crystals and final liquid lead will result. 
Thus the original lead is divided into §=-f normal crystals, assay- 
ing one-half as much as the original lead ; -| intermediary crystals, 
assaying the same as the original lead ; and % liquid lead, four times 
as rich as the original lead. The method has, however, been 
abandoned where it was tried, as it complicated the process, and as 
considerable amounts of slightly enriched leads had to be kept on 
hand. 

To carry out Pattinson's process successfully, the base bullion 
must not be very impure and a sufficient amount must be used. 
All the foreign metals contained in the lead interfere with the 
crystallization and the effectual separation of the liquid lead from 
the crystals. Ordinary lead can be sufficiently purified by poling 
(§ 109) and removing the dross that collects on the surface ; if tin, 
arsenic, -and antimony are present to any extent the lead has to be 
softened (§ 97) at a bright-red heat before the crystallization can 
proceed. Of the metals 1 commonly found in base bullion, antimony, 
bismuth, and nickel are concentrated in the liquid lead ; arsenic 
in the crystals ; copper that has not been removed with the dross 
remains equally distributed in both products. 

The kettles hold from 6 to 15 tons of lead, and the smallest per- 
missible quantity is 2J tons. 

The crystals are either taken out by a perf orated ladle or the 
liquid lead is drawn off. 

§ 92. Description of Plant and Mode of Conducting the 
Process. — The plant consists of a set of spherical kettles, from 
8 to 15 (with the method by thirds), built closely together in a row. 
Each kettle has a separate fire-place, so constructed that the flame 
shall pass beneath and behind the kettle, thence into a flue encir- 
cling it, and finally into the chimney, which has a damper to regu- 
late the draught. The details of the construction are the same as 
with the desilverizing-kettle of Parkes' process (§ 101). 

The mode of operation in outline with the method by thirds is 
as follows. In the central kettle the base bullion is melted down,, 
drossed, and poled, if necessary. The fire below is then withdrawn 
and transferred to a neighboring kettle. The cooling is promoted 
by sprinkling water on the surface from a rose. Crusts adhering 
to the sides of the kettle are pushed down into the lead, where they 
melt again. This is the work of one man, who also stirs the metal 
continuously until the smooth surface becomes rough with crystals. 
' Berg- undHuttenm. Ztg., 1889, p. 116. 



PATTINSON'S PROCESS. 



301 



His partner now inserts at the rim of the kettle a long-handled 
skimmer that has been warmed, and works it across the bottom of 
the kettle to the opposite side, then back to the middle, where, after 
jerking it to remove the liquid lead, he discharges the dry crystals 
into the neighboring kettle, generally the one to the right (" down 
the house "). The operation is continued until two-thirds of the 
contents of the kettle has been removed in the form of crystals. 
The liquid lead is then ladled into the kettle on the left ("up the 
house "). To the kettle at the right, being | full of crystals, ^ of 
lead of the same tenor is added, and the kettle at the left, being i 
full of liquid lead, is filled with a corresponding amount (f ) of 
lead of its tenor. The kettles are heated, and the cooling, crystal- 
lizing, and ladling carried on in the same way as in the original 
bullion-kettle. This becomes again filled from the crystals of the 
kettle on the left and the liquid lead of that on the right. Thus 
the operations are continued, the lead of the kettles to the right 
decreasing in tenor till that of the last one, the market-pot, assays 
from 0.3 to 0.5 ounces silver per ton ; that to the left increasing 
till the maximum of 050 ounces is reached. 

From the foregoing it will be seen that before the whole plant 
can be in working order quite a number of crystallizations have to 
be carried on, so as to have on hand the necessary amounts of lead 
of different silver-contents required to fill the kettles. 

Pattinson's process in its original form is still in use in Eng- 
land, Freiberg, and perhaps some other places. As it is improb- 
able that it will be introduced anywhere in the United States, this 
general outline will suffice. Full details are given in the works of 
Percy, 1 Kerl, 2 Stolzel, 3 Roswag, 4 Griiner, 5 and the paper by Teich- 
mann. 6 

In order to reduce the hard work necessary in withdrawing the 
crystals and ladling out the lead, as well as to insure a more regular 
crystallization and better separation of crystals and liquid lead, 
machinery was introduced into Pattinson's process; but the main 
modification of the original process, which has become the standard 
one of to-day, is that by Luce and Rozan, who stir by steam and 
draw the liquid lead off, leaving the crystals in the kettle. 

1 ''Metallurgy of Lead,'* London. 1870, p. 121. 

8 "Grundriss der Metallhiittenkunde," Leipsic, 1881, p. 225. 

3 "Metallurgie," Brunswick, 1863-1886, p. 1122. 

4 "La desargentation de plomb," Paris, 1884, pp. 211 and 267. 
6 Annales des mines, 1868, xiii., p. 379. 

6 Zeitschrift fur Berg-, Eutten-, und Salinen- Wesen in Preussen, xv., p. 40. 



302 



DESILVERIZATION OF BASE BULLION. 



§ 93. Luce and RozanV Process (Steam -Pattinson Process). 

— The advantages of the steam are, that it causes a regular crystal- 
lization' and a good separation of the lead from the crystals, and 
that it poles the lead, which being much exposed to the influence 
of the air becomes purified. Thus moderately pure base bullion, 
containing from J to | per cent, of foreign metals, can be desilver- 
ized without previous softening. It is claimed that lead with a 
little antimony and copper is even preferable, as less dross forms 
than would be the case if it were free from these metals. Of 
course, lead containing appreciable amounts of arsenic and anti- 
mony has to be softened with this process, as with any other, 
before it can be satisfactorily desilverized. 

The way in which the process is carried out at Pribram, 2 
Bohemia, may serve as an example. 

Figures 180, 181, and 183 show the general arrangement of the 




plant, consisting of two melting-pans a, one crystallizing-pot I, 
and two large conical moulds. The steam-crane is not shown; it 
is placed on the side of the crystallizing-pot, and serves to transfer 
the cakes of lead from the moulds to the storage-place, and thence 
to the melting-pans, and to tip the latter. The trough-shaped cast- 

1 Luce and Rozan, Annates des mines, 1873, iii. , p. 160; Cookson, Iron, 
September 22, 1881 ; Engineering and Mining Journal, October 8, 1881. 

2 Zdrahal, Oesterreichisches Jahrbuch, xxxiv., p. 1; and Private notes, 1890. 



PATTINSON'S PROCESS. 



303 



iron melting-pans a, each holding 1,540 pounds of lead, are placed 
behind, and 2 feet 4 inches above, the top of the crystallizer I. 
They rest with their rims on the cast-iron frame b, and are emptied 
by tipping, by means of the crane, over the inclined plate c, which 
discharges the lead through the stationary cast-iron trough d, and 




& movable sheet-iron trough (not shown), into the crystallizer I. 
Each pan has its separate fire-place (Figures 180 and 185) on the 
side, from which the gases, after passing upward (Figure 180) 
through a long flue, surround the bottom of the pan, and descend 
either directly through flue e (Figures 180, 181) to the chimney, 
or first encircle the upper part of the crystallizer (I, Figures 180, 




181), and then pass off through the flue /(Figure 181); the pas- 
sage of the gases is regulated by the dampers g (Figure 180) and 
'A (Figure 181). On the oval hearth (Figure 183) are built two 
; small walls k (Figures 180, 181, 183), in order that the flame may 
' pass close to the pan. Any lead coming from a leaking pan collects 



304 



DESILVERIZATION OF BASE BULLION. 



in the lowest part of the hearth (Figure 180), tamped with brasque, 
and is discharged outside of the brickwork. 

The crystallizer I (Figures 180, 181) is a flat-bottomed cylin- 
drical pot holding 44,100 pounds of lead, or nearly three times 
as much as one melting-pan. It has (Figure 187) near the bottom 

Fig. 183 
SECTION ON THE LINE G H. SECTION ON THE LINE J K. 

b 




two spouts, A and B, closed by slide-valves (Figures 191, 192, 193), 
for discharging the lead. At a right angle to the plane of these 
spouts is the steam-inlet C (Figures 186, 18V). The pot rests (Fig- 
ures 180, 185) on the cast-iron frame p, supported by four cast- 
iron pillars o. The top of the crystallizer (Figure 180) is covered 
with a conical hood, ending in a sheet-iron pipe, through which 
steam and dust are carried off to be condensed. The hood has 
three openings closed by doors — one at the front above the steam- 



Fig. 184: 
SECTION ON THE LINE M N. 




inlet, and two at the sides above the lead-spouts; and a small hole 
near the top, for the water inlet-pipe. The crystallizer is fired 
(Figures 180, 185) from the passage below the melting-pan a; the 
flames pass along the bottom, turn to the left, and encircle the 
lower part of the pot ; they are checked by being forced to make 



PATTWSON'S PROCESS. 



305 



their way through the narrow passage q before passing downward 
and off through the flue r. On either side of the large central 
fire-place is (Figures 184, 185) a smaller one, which serves to heat 
the discharge-spouts m before using them. Each of the discharge- 
spouts A and B (Figure 187) has a perforated cast-iron straining- 
plate to keep back the crystals when the liquid lead is being run 
off. These are held in place by wrought-iron arms b and the cast- 
iron frame c (Figure 187, 194), which is fastened by key-bolts to 
the baffle-plate d. The spouts are closed by a slide-valve (Figures 
191, 192, 193). To the flange of the spout (Figure 189) is fas- 
tened, with countersunk screws and a red-lead cement, a plate 




(Figure 190) of the same form, with one face planed smooth, 
having four openings to correspond to those of the flange; i. e., the 
central lead-discharge and the three holes near the rounded corners. 
Through these the bolts m' and m are passed, m' serving as a pivot 
for the lever o and m to tighten the guide n. To the lever is 
fastened the plate p, also having one planed face. In Figure 191 
is shown the position of the lever when the discharge is closed. In 
order to open it the nuts of the screw-bolts m are loosened, and the 
lever pushed into the second position, shown by the dotted lines. 
The lead from the crystallizer passes through the two lead-spouts 
into two tapering moulds (Figures 183, 195), each of which holds 
about 6,600 pounds of lead. The steam-inlet consists of the fol- 
lowing parts (Figure 186). On the flat bottom of the crystallizer 



306 



DESILVERIZATION OF BASE BULLION. 



are four bosses f into which fit the screws g. On the collar of 
these is placed and keyed the cast-iron circular baffle-plate d, with 
its small opening c in the centre. It serves to distribute the steam 




DETAILS OF CRYSTALLIZING KETTLE. 
Fig. 186 STEAM INLET. 



•» '' 






evenly, and to make it rise regularly in the pot. From it is sus- 
pended by an eye-bolt i with hexagonal eye, the nozzle h, into 
which is screwed the steam-pipe e. Through it passes the rod k, 
moved to and fro at one end bv the thread and cross-bar; the other 



PATTINSON'S PROCESS. 



307 



end, which is conical, fits into the conical valve-seat of the nozzle 
A, and closes or opens the steam-outlet. The steam entering at S 
(Figure 187) passes through the small annular space between pipe 
and rod, and out at h y when the valve is open. 

The mode of conducting the process is simple. Suppose the 
process to be going on and at the stage when the liquid lead has 
been drained off from the crystallizer ; the valves have been again 
closed and the crystals liquefied. One pan will be full of liquid 
lead of the same tenor in silver as the melted crystals to be dis- 
charged into the crystallizing pot, while the other will contain two 
cakes of lead that are being melted down. They will have the 
same silver-contents as the crystals remaining in the crystallizer 
after the operation to be described has taken place. The melting 
down of two cakes takes about six hours. 

The lead from the pan is run out by inserting two hooks, fas- 



Fis.193. 




tened to the chain suspended from the pulley of the crane, into the 
ears of the pan, and raising it slowly. After the lead has been 
discharged into the crystallizer the doors of the hood are closed, 
steam is introduced for two minutes, and shut off again to remove 
the pulverulent dross which has collected on the surface. Now the 
crystallization proper begins. The fire below the crystallizer is 
withdrawn and divided between the two small fire-places on either 
side, from which the lead-spouts are warmed. Steam is turned on, 
and a small jet of water is allowed to play at short intervals upon 
the surface of the lead. Every time the water is let on there are 
small explosions, and as soon as they become too violent the water 
is shut off again, while the steam enters continuously. The steam 
has forty-five pounds pressure to the square inch, and care must be 
taken to have it dry. About fifteen minutes after introducing the 
steam, the lead that has been splashed up on the upper edge of the 
pot, or on the hood, has to be removed. The steam is shut off, the 
doors in the hood are opened one after the other, and the solidified 
lead-crusts broken off with a chisel-pointed bar, and pushed back 
into the lead. This shutting off of steam to remove the lead is 



308 DESILVERIZATION OF BASE BULLION. 

repeated at least twice. While the crystallization is going on, the 
two cakes of lead required to fill again the melting-pan, just 
emptied into the crystallizer, are hoisted from below with the 
crane, and deposited one on top of the other in the pan. The 
crystallization is finished when the normal amount of steam can no 
longer overcome the resistance offered by the crystals. The result 
is that the boiling ceases, and the surface of the crystalline mass of 
lead shows only a slow, wave-like motion. Two-thirds of the 
original lead have now been converted into crystals. Water and 
steam are shut off, the slide-valves are opened, and the liquid lead 
is discharged into the moulds, which takes from eight to ten 
minutes. In these have previously been placed iron hooks, by 
which the cakes of lead, when cold, may be lifted out. The whole 
process of crystallization lasts about one hour. While the lead is 
running off, the fire from the two small fire-places is returned to 
the grate below the crystallizer, and urged in preparation for the 
next operation. The melting and other work require three hours, 
so that one operation las^ts four hours. 

In starting a series of crystallizations, 41,876 pounds of base 
bullion are taken into operation, but the dross formed soon reduces 
this amount to 39,672 pounds, of which i, or 13,224 pounds, is 
obtained as liquid lead in the two moulds. Eleven crystallizations 
are necessary to obtain market-lead from liquated base bullion 
averaging 146.12 ounces silver to the ton. The following table 
shows the average assay-value in ounces per ton of the different 
leads produced during a whole year's work. The second column 
represents the results of the same process at Eureka, 1 Nev. 



Pribram. 




Eureka. 


Market Lead, 0.43 




1.25, Market Lead. 


0.87 




2.5 


1.75 




5.0 


3.21 




9.0 


6.41 




18.0 


10.21 




30.0 


18.96 




50.0 


29.16 




75.0 


40.83 




100.0 


55.41 




150.0 


93.33 




460.0, Rich Lead. 


142.91 






Rich Lead, 262.49 






1 Curtis, ' ' Silver-Lead Deposits of 


Eureka, Nev.," monograph 


States Geological Survey, 1884, p. 


163. 





PATTINSON' S PROCESS. 



309 



Six charges are run in twenty-four hours ; two men working as 
partners attend to the crystallization, all the handling of the lead 
being done by the engineer and his helper. 

The products of the process are rich lead, desilverized lead, dross, 
and flue dust. The rich lead is cupelled, the desilverized lead is 
refined in a reverberatory furnace and moulded into market lead, 
the dross and flue dust are worked with similar products from* other 
parts of the works. The output of metal is shown by the following 
table : 



Recovered, in 


From 100 Pounds 

Base Bullion 

Charged. 


From 100 Ounces 
Silver Charged. 


From 100 Pounds 
Lead Charged. 


Rich Lead, 
Desilverized Lead, 
Scrap Lead, 
Dross, 
Flue dust, 
Loss, 


42.99 

44.76 

3.02 

9.94 

0.46 


97.36 
0.17 
0.61 
1.45 
0.07 
0.34 


42.58 
45.01 
3.03 
8.04 
0.35 
9.99 


Total, 


101.17 


100.00 


100.00 



The material consumed for desilverizing 100 tons of base bullion 
excluding the refining of the lead, is : 

Charcoal, 25 bushels. 

Bituminous Coal, 26.60 tons (for melting and desilverizing). 

" " 8.23 " (for raising steam). 

In comparing the processes of Luce-Rozan and of Pattinson, 
Cookson 1 comes to the conclusion that the former is to be preferred 
by far, as the softening of the lead is not so imperative, and the 
cost of labor only 20 per cent., and that of fuel 40 per cent., of the 
cost by Pattinson's process ; and, finally, as it produces only 33 per 
cent, the amount of drosses obtained by Pattinson. The disadvan- 
tages, greater outlay in capital, and expense of repair and renewal, 
are more than made up by the advantages. 

1 Engineering and Mining Journal, April 12, 1879. 



CHAPTEE X. 

PARKES' PROCESS. 

§ 94. Introductory Remarks. — Parkes' process is based on 
the fact that if from 1 to 2 per cent, zinc is stirred into melted 
base bullion, it will deprive the latter of its silver, and form an 
alloy, which, being less fusible than lead, and having a lower specific 
gravity, will become hard and float on the surface of the lead, 
whence it can be removed and treated separately; while the lead 
which has taken up some zinc is refined, and is then ready for the 
market. Karsten discovered in 1842 that argentiferous lead could 
be desilverized by the use of zinc, but his discovery could be 
applied in practice only when Parkes found the means (1850-1852) 
of working the zinc-silver-lead crust and refining the lead. 

The theory on which the process is based has always been that 
silver has a greater affinity for zinc than for lead, and therefore 
combines with it when added to molten bullion. According to 
Roesing 1 the statement has to be modified by saying that, while 
silver has a greater affinity for zinc than for lead, it has less affinity 
for zinc-bearing lead than either for zinc or for pure lead, and that 
this is the main cause why argentiferous lead can be desilverized 
by means of zinc. According to Alder- Wright and Thompson, 3 
zinc and silver form two definite alloys, AgZn 5 and Ag 4 Zn 5 . 

The former is an unstable compound. It has the property of 
dissolving lead, and is itself dissolved by lead to a greater extent 
than either pure zinc or the alloy Ag 4 Zn 5 . If it be kept molten for 
some- time, holding its maximum of lead in solution, it breaks up 
into Zn and Ag 4 Zn 5 , and releases some of the lead, which sinks to 
the bottom. Under the same conditions, if held in solution by 
lead, the homogeneous alloy will be divided into the above com- 
ponents, the lighter zinc rising to the surface. 

The alloy Ag 4 Zn 5 also dissolves lead, but to a smaller extent 
than a mixture of it with either AgZn 5 or free Ag would. It is, 
further, less soluble in lead than would be expected from the 

1 Zeitschrift fur Berg-, Hutten-, und Salinen-Wesen in Preussen, xxxvii., 
p. 76. 

8 Engineering and Mining Journal, December 20, 1890. 



PARKE S' PROCESS. 



311 



amount of zinc it contains. When exposed to the air it assumes a 
coppery hue. 

The low degree of solubility of Ag 4 Zn 5 in lead which must 
retain some zinc explains perhaps more definitely what Roesing 
called the small affinity of silver for zinc-bearing lead. 

Before zinc added to the base bullion can take up any quantity 
of silver, it combines with the gold and copper contained in the 
lead and saturates this, the amount taken up depending on the 
temperature of the lead (§ 11,*). By the use of zinc a market 
lead very low in copper is thus obtained, and by successive addi- 
tions of zinc very small amounts of gold can be concentrated in 
a separate crust (the gold or copper crust) with some silver, and 
extracted at a profit. There is a difference of opinion ' as to 
whether gold or copper combines first with the zinc. It would 
seem that it must be gold, as no desilverized lead is free from copper, 
but it never retains the least trace of gold. 

In order to desilverize argentiferous lead with zinc successfully, 
it is necessary that the lead and zinc be reasonably pure. Tests 
made by Kirchhoff 2 on base bullion containing 4.5 percent, foreign 
metals, such as copper, arsenic, antimony, bismuth, and zinc, showed 
that 2.87 per cent, zinc was required to desilverize the lead when 
the bullion had not been softened, while 1.75 per cent, was suffi- 
cient if softening had preceded the desilverization; the relative 
quantities of market lead produced were 43 and 72 per cent, of 
the bullion charged. The following table shows how the silver- 
contents decreased with each addition of zinc: 



Number of additions. 


Not Softened. 


Softened. 


Ounces silver 
per ton. 


Pounds of 
zinc. 


Ounces silver 
per ton. 


Pounds of 
zinc. 


After drossing. . . 
1 


85.60 

85.50 

85.30 

83.80 

83.50 

83.00 

48.20 

8.20 

0.80 

0.15 


250 
250 
150 
100 
100 
100 
100 
70 
30 


94.90 
85.60 
47.60 
16.10 
1.70 
0.18 


150 
150 
150 
150 
100 


2 


3 


4 


5 


6 


7 


8 


9 





1 Percy, "Metallurgy of Lead." p. 174. 

2 Metallurgical Review, vol. i., p. 224; or Dingier, Polytechnisckes Journal, 
vol. ccxxviii., p. 265. 



312 DESILVERIZATION OF BASE BULLION. 

With the crude lead the first five zinc additions served only to 
remove the impurities to such a degree that the desilverization 
could begin. That the first addition of zinc to the softened lead 
took up so little silver shows that the lead must have been verv 
coppery. 

Of the three metals, copper, arsenic, and antimony, that 
principally interfere with desilverization, antimony is the least 
objectionable, as lead with as much as 0. 7 per cent, of antimony, 
assaying 41 ounces silver to the ton, is desilverized without being 
softened, 1.3 per cent, of zinc being required, and 81.34 per cent, 
of market lead being produced. 

Arsenic not only retards the desilverization greatly, but seems 
to prevent the zinc crust from separating satisfactorily from the 
lead. In skimming a kettle the usual beautiful, smooth, dark-blue 
surface is not seen, but it shows instead a rough, grayish-white 
surface ; and even if skimming be continued until the lead solidifies, 
the surface will hardly change in appearance. 

That copper combines with zinc before silver has already been 
mentioned. 

Of the other two metals contained in Kirchhoff's bullion, zinc 
has a favorable effect; bismuth is indifferent, as it remains for the 
most part alloyed with the lead. 

The zinc used must be pure, if the desilverization is to proceed 
satisfactorily. In experimenting with cheap zinc containing iron, 
obtained from galvanizing works, the writer found that the process 
was so retarded, and the quantity of impure zinc required so great, 
that no saving at all was effected by the use of the inferior material. 
Jernegan ' records a similar experience, and Fohr 2 mentions that he 
required four times the usual amount of zinc, which was found to 
be due to its impurity. It contained 2.75 per cent, lead, 0.61 iron, 
0.077 copper, and traces of tin, arsenic, antimony, cadmium, sul- 
phur, and carbon ; in good brands the iron appears only in the 
second decimal. 

From the effect that foreign metals have on the result of the 
desilverization, it is clear that all argentiferous lead that contains 
them to any extent has to be softened. All or almost all Ameri- 
can zinc-desilverizing works buy base bullion and ores in the open 
market to a greater or less extent. The bullion is therefore always 
liable to contain some arsenic or antimony, and every refinery 
softens the lead before it attempts to desilverize. 

1 Trans. A. I. M. E., ii., p. 288. 3 Berg- und Huttenm. Ztg., 1888, p. 28. 



PARKE 'S' PROCESS. 313 

§ 95. Outline of Plant and Process.— The general plan of a 
desilverizing plant varies somewhat according to its location and 
the practice that prevails. All the arrangements, however, must 
be such as to require as little handling of lead and by-products as 
possible. In fact, the bulk of the lead, when charged into the 
softening-furnace, will in a modern plant never be handled again 
until it is ready for shipment. The result is that a section through 
a refinery will show the form of a terrace. 

The general arrangement of a refining plant is given in plan and 
section in Figures 196 and 197. On the highest level are the soft- 
ening-furnaces, which receive the base bullion and prepare it for 
the kettles. The latter are a stage lower, and there the softened 
bullion is desilverized. The apparatus for liquating zinc-crusts is 
also upon the same floor. In the drawing, each desilverizing-kettle 
has close to it only one liquating-kettle, with a small kettle for 
liquated lead, showing that no distinction is made between gold 
and silver crusts. The liquated crusts pass from this floor into an 
adjoining building, placed to the right or the left of the main 
building, the floor of which is on a level with the scale-floor. It 
contains two departments — the retort-room, where the crusts are 
distilled, and the cupelling-room, where the retort-bullion is turned 
into silver, or dore silver bars. In the plant shown by the figure 
only dore bars will result. Following the desilverized lead, the 
refining-furnaces are reached on the next level, in which the 
desilverized lead is dezincified. Thence it passes into the merchant- 
kettles, and from these into the moulds placed in the lead-pit. 
The market lead is loaded on trucks on the scale-floor, that are 
of the same construction as the bullion-receiving trucks ; they 
are run on scales, and the weighed lead is transferred into the 
cars on the loading-track. The plant for working the by-prod- 
ucts will be placed on the side of the main building, opposite to 
where the zinc-crusts are treated. The manner of dealing with 
these varies greatly in different refineries, and will be discussed 
later on. 

Roesing ' draws a comparison between the arrangement of a 
refinery on a horizontal plane and that on an inclined plane, which, 
on the whole, is not favorable to the latter. 

According to the outline given, Parkes' process is best treated 
under the following heads : 

1 Zeitschrift fur Berg-, Hutten-, und Salinen-Wesen in Preussen, xxxvi., p. 
103; Berg- und Huttenm. Ztg., 1888, p. 337. 



314 DESILVERIZATION OF BASE BULLION. 

Receiving Base Bullion. 

Softening Base Bullion. 

Desilverizing Softened Bullion. 

Refining Desilverized Lead. 

Moulding Refined Lead. 

Treatment of Zinc-Crusts. 

Treatment of By-Products. 

Table of Desilverization. 

Conclusion. 

Comparison between Pattinson's and Parkes' Processes. 

§ 96. Receiving Base Bullion. 

Beceivmg Base Bullion. — The base bullion arrives at the 
refinery in car-loads of from ten to fourteen tons. The track 
(receiving-track of Figure 196) is laid so low that the bottom of the 
car is on a level with the " upper platform " of the works. Along 
the whole length of this platform, and parallel with the railroad, 
runs a narrow-gauge track, from sixteen to twenty-two inches, 
which bears a number of strongly-built low bullion-trucks. They 
may be built as follows. A frame consisting of two pieces of 
channel-bar iron, three feet long, is fastened to the two axles of 
the wheels, and steadied by two iron bands running diagonally. 
The bullion is carried out from the car and loaded upon the truck 
standing before the car-door ; when this is filled to a height con- 
venient for lifting (about 3^ feet), it is moved on, and another takes 
its place. The trucks are run on scales placed at one or two points 
in the "bullion and scale-shed," the bullion is weighed, and is 
then sampled from one truck directly upon another, which then 
moves straight to the softening-furnaces, or to places near them, 
and no more handling is required before the bullion is charged 
into the furnaces. The bullion produced in the smelting depart- 
ment of the refinery is loaded at the blast-furnaces or reverberatory 
furnaces, on the same kind of truck, brought by an elevator to 
the "bullion and scale-shed," and then passes to the softening- 
furnaces. 

§ 97. Softening Base Bullion. 

Introductory Bemarks. — The object of softening is to separate 
from the base bullion produced in the blast-furnace, impurities, 
such as copper, sulphur, tin, arsenic, antimony, etc., that would 
interfere with the desilverization. It comprises two processes, 
liquation and oxidation. By the former, metals and their com- 



r 






! * ! 

is : 

i t- ; 

! o ! 

! z ' 
i > | 

! u i 

O i 

UJ , 

i a 

[_A_L_ 


s 

DC 

o 

u_ 

< 

_l 

CL 

c 

111 

Q. 
Q- 

3 


:- 


i ! 




- 



: ■. 



H 



! L 



HOC . 










tfOT-C 




PLAN OF DESILVERIZING PLANT, WITH 30-TON KETTLES. 
.Fi a . LOO. 









.3 AHOIT033 






F 



03 M 













TWTAJW Okiismavjieaa ^o hajs 



PARKE S' PROCESS. 



315 



pounds held in solution by the red-hot blast-furnace-lead are sep- 
arated out again from the readily fusible lead by melting it down 
slowly at a low temperature. By the latter, metals alloyed with 
the lead, and more easily oxidized than the lead, are removed by 
heatin°- it to a bright-red heat with access of air, with the result 
that these metals are converted into oxides, and, combining with 
the lead oxide, are drawn off either as a powder or a slag from the 
surface of the metallic lead. 

When base bullion is melted down slowly in a softening- 
furnace at a low heat, there rises to the surface a dark-colored, 
half-melted, pasty substance, the furnace-dross, consisting of a 
mixture of lead, copper, sulphur, arsenic, etc. The slower the 
meltino- down, the more effectual will be the separation of the 
copper from the lead. The following analyses show the purifica- 
tion effected in the lead by liquating and the composition of the 
dross when it has been freed as much as possible from adhering 
lead : 





Clausthal. 1 


Lautenthal. 2 


Freiberg. 3 


Before 


After 


Before 


After 


Before 


Liquated 




Drossing. 


Drossing. 


Drossing. 


Drossing. 


Drossing. 


Dross (5jJ). 


Pb 


98.92944 


99.0239 


98.96475 


99.1883 


96.667 


62.40 


Cu 


0.1862 


0.1096 


0.2838 


0.0907 


0.940 


17.97 


Cd 


trace 


none 


trace 


none 






Bi 


0.0048 


0.0050 


0.0082 


0.0083 


0.066 


none 


Ag 


0.1412 


0.1420 


0.1413 


0.1440 


0.544 


0.17 


a! 


0.0064 


0.0053 


0.0074 


0.0032 


0.449 


2.32 


Sb 


0.7203 


0.7066 


0.5743 


0.5554 


0.820 


0.98 


Sn 


none 


none 


none 


none 


0.210 


0.04 


Fe 


0.0064 


0.0042 


0.0089 


0.0048 


O.027 


0.43 


Zn 


0.0028 


0.0017 


0.0024 


0.0015 


0.022 


0.07 


Ni 


0.0023 


0.0017 


0.0068 


0.0038 


I 0.055 


1.09 


Co 


0.00016 


trace 


0.00035 


trace 


S 






.... 




0.200 


4.00 















1.87 


Slag, as] 


l, heart h-m< 


iterial .... 








8.66 











The analyses from Clausthal and Lautenthal demonstrate that 
the character of a comparatively pure lead is improved by melting 
down slowly and drossing. By comparing the two Freiberg 
analyses the degree will be seen to which the foreign matter of a 
very impure base bullion may be removed ; viz., nearly all the 

i Hampe, Zeitschrift fur Berg-, Hutten-, und Salinen-Wesen in Preussen, 
xviii., p. 203. 
3 Ibid. 3 Schertel, Berg- und Huttenm. Ztg., 1882, p. 293. 



316 



DES1LVERIZATI0N OF BASE BULLION. 



sulphur, 93 per cent, copper, 96 per cent, nickel and cobalt, 25 per 
cent, arsenic, and only 5.8 per cent, antimony, and 1.54 per cent, 
silver. • Bismuth remained entirely in the liquated lead, and all the 
tin excepting 0.9 per cent. 

On melting down this dross in a crucible, Schertel J obtained a 
product consisting of lead, speise, and matte, in well separated layers. 





Lead. 


Speise. 


Matte. 


As 


0.34 
1.79 
96.50 
0.08 
0.75 


37*60 

25.68 

8.60 

27.00 


47.70 

32.80 

0.25 

1.15 

17.72 


of;::;;:;:;::::::;;:.;:::::: 


Pb 


Ni 


As 


S 





The absence of iron in either speise or matte proves that these 
impurities do not result from finely-divided blast-furnace speise 
or matte being dissolved by the lead, as has been often thought. 
It tends to show that, being held in solution by the metal, the 
impurities unite on liquating to form a compound that is not fusi- 
ble at the temperature at which the lead was melted down, and that 
the concentration of copper in dross is due probably to the presence 
of sulphur and arsenic and not to the separation of an alloy of lead 
and copper. 

If a sample be taken of the lead after drossing, poured into a 
small mould, and allowed to cool slowly, a crystalline, bright, 
pewter-white spot will appear on the slightly depressed, dull, gray- 
ish-white surface, which, in addition to the hardness of the lead, is 
characteristic for the presence of arsenic and antimony. 

If the temperature be raised to a good red heat and the air per- 
mitted free access, the impurities contained in the lead will oxidize 
one after another : first tin, then arsenic and antimony. The surface 
at first will become quickly covered with dark-yellow skimmings, 
which vary from powdery to pasty, but are not fused on account 
of the tin ; these are called tin-skimmings. They consist mainly of 
antimoniate and stannate of lead and antimoniate of tin, and are 
worked by themselves (§ 120). As soon as the tin-skimmings have 
been drawn off from the surface of the lead, this begins to give off 
fumes of arsenic and antimony, and arseniate and antimoniate of 
lead begin to form ; the former is a light brown, the latter dark 
brown to black ; both are fused and are drawn off together as anti- 

1 Loc. cit. 



PARKE S' PROCESS. 317 

mony-skimmings when the furnace has been sufficiently cooled down 
for them to solidify. Towards the end of the operation the anti- 
mony in the skimmings will be replaced by lead until the black 
color has changed to the greenish-yellow of litharge. 

Samples are then taken to see how far the softening has pro- 
gressed. Before the antimony has been removed, a sample of the 
bullion taken in a ladle will "work/' i.e., small particles of melted 
black skimmings will float on the surface of the lead, with a rotary 
motion which resembles that of particles of grease on hot water. 
As the softening approaches the finishing-point, the globules be- 
come less in number and smaller in size, a thin coating of yellow 
litharge forms more readily on the red-hot lead, and finally no 
more globules are seen and litharge forms quickly. 

When a sample of the lead is poured into a mould, allowed to 
cool slowly, skimmed with a flat piece of wood, it will, when it has 
solidified, have lost the characteristics of arsenic and antimony, 
and the surface of the bar will have assumed a rich indigo-blue 
color ; the ease with which it can be scratched with the finger-nail, 
and the lustre on a freshly-made incision, show the change that has 
taken place. It can also be tested by cupelling a small sample ; if 
any incrustation is left on the cupel, the lead is not sufficiently 
softened. 

§ 98. Furnaces. — The reverberatory is the furnace universally 
used in this country to soften base bullion. In some European 
works the liquation is carried on separately from the oxidizing 
smelting. The furnace is a real liquating-furnace with an inclined 
hearth, from which the lead runs off into an outside kettle, whence 
it is ladled out ready for the softening-furnace proper. The dross 
obtained is as free from lead as liquation can make it (see Freiberg 
analysis, ante). It is, however, more economical to make the entire 
softening process a continuous one, as it reduces the apparatus and 
the number of men necessary. Therefore in American refineries, 
both liquating and oxidizing smelting are carried on in the same 
kind of reverberatory furnace, and the drosses which contain con- 
siderable lead are all liquated in a single liquation-furnace, which 
is always kept running, and serves for the whole plant. 

In some works the bullion is melted down in the desilverizing- 
kettle, drossed, heated to bright redness, and oxidized by introduc- 
ing dry steam, which continually renews the surface of the lead, 
and thus hastens the elimination of arsenic and antimony. 

This is the most expensive way. It consumes much fuel, is 



318 DESILVERIZATION OF BASE BULLION. 

lengthy, forms a very large amount of oxidized product, and ruins 
the kettle. At the level, where the antimoniate of lead is in con- 
tact with the iron, it is eaten out. Even a kettle made especially 
thick at this place lasts only a short time. 

The reverberatory furnaces used for softening are generally 
large enough to hold from 8 to 10 per cent, more bullion than 
the kettle into which they discharge their contents. About ten 
years ago furnaces were constructed to hold from 15 to 20 tons of 
base bullion ; their size has been gradually increased until most 
furnaces now hold from 33 to 35 tons of base bullion, furnishing 
30 tons of softened bullion to the kettle. At some works, furnaces 
holding 50 tons are in successful operation, but they form the 
exception. 

The construction of the different furnaces is the same in many 
points. The hearth is elliptical or rectangular in plan ; the length 
being to the width as lj : 1, or often as 2 : 1. It is built of fire- 
brick, enclosed by an iron pan to prevent the leakage of lead. In 
section it is dish-shaped ; it is shallow, its depth varying from 1 1 to 
16 inches ; in exceptional cases it reaches 22 inches. Its slope 
depends upon whether the tap-hole is located on the side or the 
flue end of the furnace, being from 2 to 5 inches. 

As regards detail, there is considerable variety in construction. 
The pan which is to hold the hearth used to be made of cast iron ; 
it rested, on transverse rails supported by brick walls running 
longitudinally. Thus the bottom of the hearth was effectually 
cooled by the air circulating beneath the pan. In order to relieve 
it from strains, the pan was allowed to stand free, the skewbacks 
supporting the roof being separate heavy castings, held in place 
by buckstays and tie-rods. Notwithstanding all these precautions, 
a cast-iron pan generally cracked after it had been in use for a 
little while. To-day there may still be found a few old furnaces 
with cast-iron pans ; a new furnace, however, will always be built 
into a wrought-iron pan. The softening of lead that contains a 
few per cent, of antimony often takes considerable time and 
requires a pretty high temperature. Thus the brick suffers greatly 
by the corroding action of the antimoniate of lead and the litharge 
forming on the surface of the lead. The best fire-brick soon 
begins to be eaten out if much hard lead comes to the furnace, and 
patching with a mixture of raw and burnt fire-clay, or raw clay 
and coke, after every few charges, has to be resorted to in order to 
preserve the side walls. As the furnace has to be somewhat cool 



PARKE S' PROCESS. 319 

to make this repairing effective, much time is wasted. An improve- 
ment was made by introducing a 2-inch pipe between the two 
courses of fire-brick forming the side-lining of the furnace, and 
allowing water to circulate through it. The inside course was 
eaten away by the litharge to a thickness of from 2 to 3 inches as 
quickly as before, but then the corrosion proceeded only very 
slowly, and the life of the side-walls was thus greatly prolonged. 
With this encouragement the water-cooling has been carried to the 
extreme of enclosing the wrought-iron pan holding the hearth in 
another one, leaving 3 or 4 inches space between the two, in which 
water circulates, thus cooling not only the sides but also the 
bottom of the furnace. The larger number of softening-furnaces 
in use to-day have two wrought-iron pans with water circulating 
between them. The outer pan is supported in the same manner 
as the former cast-iron pan ; the inner pan rests also on rails which 
are laid in the same direction as the walls ; stay-bolts connect the 
outer and inner pans. While without doubt this mode of cooling 
is very effective, there is unquestionably too much of it, consider- 
ing the amount of fuel that is required to keep up the temperature 
necessary to soften the lead within the given time. Air-cooling 
alone has always been sufficient for the bottom ; water-cooling is 
necessary for the sides only. 

At several works the latest softening-furnaces have water 
jackets only at the sides, and one of these, given in Figures 198- 
204, is chosen to illustrate some of the details. The side eleva- 
tion (Figure 198) shows the fire-place a and the hearth inclosed by 
the water jacket b and the pan c. Between fire-box and hearth is left 
an air space d (Figures 198, 199, 202). To counteract the bulging 
out of the pan and bending of the jacket, due to the expansion of 
the hearth, 3-inch I-beams are placed horizontally behind the jacket, 
and 7-inch I-beams behind the pan. They are not shown in the 
drawing. The discharge of the lead takes place at the flue end 
of the furnace through the spout e. The products of combustion 
pass off through the roof at/* (Figure 202) into the horizontal flue 
g (Figure 200), which crosses the furnace, and then, passing down 
vertically close to it, leads into the main canal underground. The 
bullion is charged through two doors h on one side of the fur- 
nace (Figure 199); the drossing and skimming takes place on the 
opposite side, through three doors i (Figures 198, 199). The charg- 
ing doors are on a level with the upper edge of the water jacket ; 
the drossing and skimming doors are let in 3 inches. The pan of 



320 DESILVERIZATION OF BASE BULLION. 

the hearth rests on 7-inch I-beams j 9 placed transversely, and these 
rest on two brick walls k running longitudinally. The hearth (Fig- 
ure 201) is inclosed up to the level of the charging doors, i.e., to 
the depth of 42 inches, by an iron pan made of -§-inch boiler iron. 
To the upper part of the pan is attached the water jacket, consist- 
ing of a boiler plate $ of an inch thick and 21 inches wide, and 
two pieces of bar iron m 3 inches square, riveted to the pan. 
Stay-bolts, 9 inches apart, connect the two plates. The jacket has 
a lj-inch water inlet and outlet pipe. To insure the complete fill- 
ing of the jacket, small pieces of pipe pass through the upper 3-inch 
bar. The jacket has further hand-holes to clean out at intervals 
the mud that settles out from the cooling-water. These details are 
not shown in the figures. 

Another side jacket that is found at some furnaces has a some- 
what different construction. It is open at the top and covers the 
entire side of the pan. Near the bottom it is slightly bent, and 
one row of rivets joins the bottom and side of the pan with the plate 
forming the jacket. Both kinds of jackets are in satisfactory use. 

The manner followed in putting in the hearth (Figure 201) 
varies somewhat at different works. Usually a layer of brasque is 
first carefully tamped in, and then so cut out that the course of 
brick, laid endwise upon it, shall bring it into the desired shape 
and give it the necessary inclination toward the tap-hole. Assum- 
ing this to be at the flue end of the furnace, as is most common, 
the thickness of the brasque there will rarely exceed 2 inches, 
increasing to 5 inches at the centre of the fire-bridge. In the 
figure very little brasque n is observed in the centre ; it increases 
to a thickness of 7 inches at the sides. It is made so thin because 
there are two courses of brick instead of one, as is usual. The 
lower course o is an inferior grade of fire-brick, set dry, and grouted 
with a mixture of clay and cement, so as to stop up the joints and 
prevent any percolation of lead. The upper layer p is maVie of 
the best fire-brick available. In putting down the bottom the 
bricks have to be joined as tightly as possible. For this purpose 
they must first be carefully selected and fitted by rubbing them 
together until all roughness is removed. Each brick is dipped into 
water and then into a clay-mortar having the consistency of very 
thin gruel, then put in place and driven with the hammer against 
the brick it is to face. This makes the joint as close as possible 
and prevents the passage of lead. The sides of the furnace are 
built with the same care as the bottom. Commonly they rest on 



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I DETAIL OF 
u THE TAP 



PARKED PROCESS. 321 

the curved working-bottom to prevent this from rising. In Figure 
201 the sides q rest on the lower course, while the working-bottom 
p abuts upon them. Contrary to expectation, the bottom shows 
no tendency to rise. The reason for putting in the sides in this 
way is their convenience for repairing when eaten out. The corro- 
sion and the lead do not, with a good bottom, reach the lower 
course of brick. When the side-wall has been torn out till sound 
bricks are reached, it can be built up again on a solid, smooth, hori- 
zontal layer, instead of having to be placed upon the protruding 
headers of the working-bottom, The roof is supported on either 
side by two rails s serving as skewbacks. Usually it has a slope 
from the bridge to the flue, and is not horizontal, as represented by 
Figure 202. The furnace is bound in the usual way with buck- 
stays (old rails) and tie-rods. The means of tapping the furnace 
deserves special mention. In Figure 200 the water jacket is seen 
to inclose the tapping-opening. Details are shown in Figures 
203 and 204. The 2-inch tap-hole t is a conical opening in the 
cast-iron plate >/, which tills the open space between the outer and 
inner plates of the jacket. Two bolts v pass through the jacket 
and plate, and then through the flanges of the spout placed on the 
outside, where they are tightened with nuts. Commonly the tap- 
hole is closed by simply ramming a clay plug into it. To make 
the breaking away of the clay impossible, the tap-hole in the figure 
is closed by an iron plug coated with clay. It is held in place 
by an iron wedge driven vertically between it and a horizontal 
piece of flat iron, held in place by the vertical ribs ic on either side 
of the spout. 

Finally, as to the firing of the furnace. The fuel commonly 
used is bituminous coal; at some works this has been replaced by 
crude oil. With a good grade of bituminous coal, natural draught 
is suflicient to soften the lead in the required time. Many works 
use slack coal, and then blast under the grate becomes necessary. 
At some works a series of small blast-pipes is introduced through 
the roof, above the fire-bridge, with very good effect. The writer 
recalls one instance where, with the use of blast both under the 
grate and in the roof, and an impure bituminous coal, Colorado 
base bullion of ordinary hardness was softened in 50-ton charges in 6 
hours. This is probably the best work on record. The good effect 
of cold blast in the roof would suggest the admission of hot air, as is 
done in Lake Superior copper-refining furnaces, 1 where a flue as- 
1 Egleston, Trans. A. I. 31. E., ix., p. 690, plate ii. 



322 DE8ILVERIZATI0N OF BASE BULLION. 

cending in the side wall of the fire-place passes through the roof, and 
enters the furnace just above the bridge, delivering there the heated 
air to effect complete combustion. If necessary, additional heated 
air might enter the furnace through openings in the fire-bridge. 

In the use of oil as fuel, air under a pressure of 8 or 9 ounces 
has begun to replace the different forms of Korting inspirators, 
although a larger number of burners is required to furnish the 
necessary amount of oil. 

§ 99. Mode of Conducting the Process. — The mode of oper- 
ating the softening-furnace is about as follows. The buHion is 
charged through the two charging-doors by means of a long- 
handled paddle, and melted down slowly. 

The paddle is a rectangular iron bar about 8 feet long, made of 
lj-inch iron ; one end is flattened out for a distance of 2 feet 6 
inches to the width of 3 inches, to receive a bar of bullion, while 
the other is rounded off and bent to a ring. A lug is often cast on 
either side of the door-frame to support a roller. It serves as 
bearing for the paddle instead of the door-frame, and thus facilitates 
the manipulation. 

In some works the bullion is all charged at once ; in others, 
where the charging, softening, desilverizing, refining, and mould- 
ing are given in contract to a crew of four men, the bullion is 
charged at intervals when the kettle and refining-furnace do not 
require their attention. The softening of the base bullion, as well 
as the refining of the desilverized lead, is regulated by the desilver- 
ization which takes the longest time (from 18 to 20 hours for a 30- 
ton kettle). When a kettle of bullion has been desilverized, the 
refining-furnace (average time required 12 hours) must be empty 
to receive the desilverized lead, and the softening-furnace (average 
time required 14 hours) ready to furnish softened bullion to the 
kettle. Upon this general scheme the whole of the refinery must 
be based to be properly continuous. 

The bullion is melted an hour or more after it has been charged. 
It is stirred to detach some of the lead held in suspension by the 
dross ; sometimes fine coal is spread over it and stirred in. This is 
very effective when the bullion is pure, so that little dross rises to 
the surface. With impure bullion considerable fuel is required to 
have any effect, and there is danger of the temperature becoming 
too high and the lead taking up again some of the impurities that 
had separated out. The dross is removed by a rabble or a rec- 
tangular skimmer. 



PARKES* PROCESS. 323 

The head of the rabble is made of |-inch iron, and is 3 by 12 
inches, the handle of -J-inch iron, and 10 feet long. The handle of 
the skimmer is of the same length and thickness as that of the 
paddle ; the perforated part is made of J-inch iron, and is 10 by 12 
inches, the perforations being -§ inch in diameter. 

With both tools about the same amount of lead is withdrawn 
with the dross from the furnace, so there is little choice between 
them : some prefer one, some the other. With either, the handle 
often rests in a hook suspended by a chain from the roof, thus 
facilitating the work. The operator removes from one side the 
dross, which his helper on the other side collects with a rabble, 
pushing it toward the door or upon the skimmer. The dross, while 
being removed from the furnace, is collected either in a slightly 
conical cast-iron mould running on wheels, e.g., 2 by 3 feet at the 
base and 14 inches deep, made of f-inch iron, or preferably in an 
iron two-wheel barrow with perforated bottom, in order that some 
of the lead carried out with the dross may run off on the cast-iron 
plate in front of the skimming doors and be returned to the 
furnace. 

The dross drawn off is weighed, and a sample is taken from the 
lead remaining in the furnace to be assayed. The weight and 
assay-value of the bullion charged being known, the weight of the 
dross and the assay of the residual hard lead give the data neces- 
sary to calculate the total silver contained in the dross, and with it 
its assay. The amount of dross formed is about 4 per cent, of the 
bullion charged, and assays about 80 per cent, of lead. It is freed 
from some of its lead in a liquating-furnace, and will then have a 
composition similar to that of the Freiberg analysis (see ante). 

The tin-skimmings that form on raising the temperature after 
drossing are removed in the same way as the dross. 

With the antimony -skimmings it is customary to cool the fur- 
nace by throwing open the doors, in order that the antimoniate of 
lead floating on the surface may harden, and be then taken off in 
the form of a thin crust. If the bullion is very hard, skimming once 
will not be sufficient to soften it. The furnace is therefore heated 
up again, and as soon as the surface of the lead is well covered, 
the cooling and skimming are repeated ; ordinarily two operations 
are sufficient, but sometimes three are necessary. To hasten the 
cooling of the furnace, slacked lime is sometimes spread over the 
metal bath. Some refiners add lime to the furnace after drossing, 
with the idea that an antimoniate of lime is formed, and thus less 



324 DESILVERIZATION OF BASE BULLION. 

lead is oxidized during the softening. This effect of lime still 
remains to be proved. Any addition of lime to the furnace has 
the great disadvantage of interfering with the subsequent liquation 
of the antimony-skimmings, and is therefore better dispensed with 
altogether. 

If the bullion is very hard, the addition of litharge from the 
cupelling furnace greatly shortens the time required for soft- 
ening. 

Another method of hastening the softening is the introduction 
of dry steam to stir the lead, thus continually exposing a fresh 
surface to the oxidizing action of the air. This is done by intro- 
ducing through each of the charging doors a 1-inch pipe, to the 
end of which are screwed, by means of a T, two pipes having a 
number of perforations on either side and closed at the ends. The 
main pipe is bent so that when it is introduced into the furnace 
and held in place by the closed furnace door, which has been 
weighted, the two pipes at its ends will be pressed down into the 
lead, and run parallel to the sides of the furnace. While the 
introduction of steam may shorten by one-third the time required 
for softening, it has the disadvantage that it forms a large amount 
of skimmings, and that the swash of the lead oxide and antimoniate 
strongly corrodes the sides of the furnace. It is, therefore, to be 
used only in extreme cases. 

A third method to be mentioned is the one in use at Freiberg, 
where bullion rich in tin, arsenic, and antimony is softened. 
Blast is introduced on either side of the fire-bridge, and the skim- 
mings are removed at the flue end of the furnace as fast as thev 
form. The tool used is a long iron hook, to which is fastened 
a triangular piece of wood, say 8 inches long. With it the skim- 
minors are drawn out of the furnace in a thin stream. In order to 
facilitate the work, and to enable the workmen to pass gently over 
the surface, and thus remove only skimmings, but no lead, the 
handle is supported by a hook suspended from the roof. 

After the last skimmings have been removed, the doors are 
thrown open to cool the lead before it is tapped into the kettle. 
The skimmings are weighed and a sample of the softened bullion 
in the furnace is taken for assay ; thus the silver contained in the 
skimmings can be calculated. The amount of skimmings found is 
about 5 per cent, of the weight of the bullion charged. The fuel 
consumed for the entire softening is about 156 pounds of soft coal 
per ton of bullion charged. 



PAMKES' PROCESS. 325 

A sample of flue dust from the softening-furnace assayed 15.9 
per cent, lead, 9.1 ounces silver, and 0.16 ounces gold. 

A new furnace bottom absorbs a considerable amount of base 
bullion for which it is difficult to give any figure. One peculiarity 
still needs a satisfactory explanation, namely, that a larger propor- 
tion of gold collects than of silver, considering the average com- 
position of the bullion treated. 

§ ioo. Desilverizing Softened Bullion, 

Introductory -Remarks. — From the softening-furnace the lead, 
when sufficiently cool, is tapped into the desilverizing-kettle, 
which has been whitewashed with lime water and heated to the 
point where a splinter of dry wood thrown on the bottom will 
ignite readily. The whitewashing facilitates the removal of silver- 
crusts which adhere to the sides when the kettle is cooling. If 
the lead were tapped into a cold kettle, this would be liable to 
crack on the bottom, and the time for bringing the lead up to the 
required temperature would be unnecessarily prolonged. The lead 
runs into a trough of cast iron, -J inch thick, placed beneath the 
discharge-spout of the furnace. In order to decrease the amount 
of dross, the lead runs from the trough into a cast-iron pipe placed 
upright in the kettle. 

The kettle-dross formed amounts to about one per cent, of the 
bullion charged. It is skimmed off and added to the next charge 
in the softening-furnace after the furnace-dross has been taken off. 
The kettle is now ready for the addition of zinc. 

The quantity of zinc necessary varies according to the purity 
of the lead, and increases on the whole with the amount of silver 
present. Roswag's ' new formula is 

Z = 10.39 + 0.035 T, 

where Z = kilogrammes zinc to be added to one metric ton of lead, 
and T = grammes silver in 100 kilogrammes lead. 
This corresponds to 

Z' = 23.32 + 0.223 T", 

where Z = pounds zinc to be added to 2,000 pounds of lead and 
T' = ounces silver per ton. 

Roswag also formulated the quantities for zinc Illing a found 

1 " La desargentation de plomb," Paris, 1884, p. 241. 

2 Ztschrft. f. B. H. u. S. W. i. P., xvi., p. 51. 



326 



DESILVERIZATION OF BASE BULLION. 



necessary to desilverize lead of varying tenor in silver. They 

are : 

Z" kilogrammes =11.66 + 0.0325 T" 

(grammes silver in 100 kilogrammes lead) 

for every metric ton of base bullion ; which corresponds to 

Z'" pounds = 20.78 + 0.24 T" (ounces silver per ton) 

for every 2,000 pounds of base bullion. 

These general formulae probably give an approximate idea of 
the total amount required to desilverize base bullion running low 
in silver, say 30 ounces. For rich bullion the figures are too high. 
How the amount of zinc to be added increases with the silver-con- 
tents, irrespective of the zinc recovered later by distillation, is shown 
in the following figures given by Plattner. 1 



Assay of base bullion, 
ounces per ton. . 


Per cent, of zinc added. 


28.09 
111.56 
148.16 
245.00 


1.34 

1.84 
1.96 
2.45 



In practice it has so far not been found possible to desilverize 
a rich base bullion with a single addition of zinc. The richer the 
bullion the less difference is there between the assay-values of the 
zinc-silver crust and the residual lead. From three to four zincings 
are therefore necessary. Low-grade bullion can be and is de- 
silverized by two zincings. The aim in desilverizing must be to 
concentrate as much silver as possible into one zinc-crust, so as to 
utilize all the power of the zinc. This is best done by first adding 
sufficient zinc to remove the gold and the copper (as gold-crust) 
with as little silver as possible, and to saturate the lead ; then the 
bulk of the zinc is added, which takes up enough silver to form the 
rich zinc-silver-lead alloy (the first silver crust). One or two sub- 
sequent additions of zinc will completely desilverize the lead, but 
the zinc contained in these crusts will, under suitable conditions, 
combine with more silver ; hence the second and third silver-crusts 
are added as fresh zinc for the first desilverization, and two-thirds 
of the zinc contained in them is available as new zinc. The crust 
obtained from a kettle assaying 60 ounces or less per ton always 
1 Berg- und. Euttenm. Ztg., 1889, p. 117. 



PABKES' PROCESS. 



327 



goes back to the kettles. The second silver-crust ranges ordinarily 
from 18 to 30 ounces, the third silver-crust from 0.3 to 0.5 ounces 
per ton, while the first never ought to run lower than 2,000 ounces. 
The following analyses show the composition of liquated first 
zinc-crusts obtained from low-grade bullion that has been drossed 
only and still retained the antimony. 





Altenau. 1 


Lauten- 
thal.2 


Mechernich. 3 


Friedrichshiitte. 4 


Pb 

Zn 

Cn 

Ag 

As 

Sb 

Cd 

Ni 

Fe 

PbO . . . 
ZnO.... 
Bi 2 3 . . 
H 2 3 ... 
Fe 2 3 . . 


1880. 


1880. 


1875. 3 


1884.* 


Average. 
1884." 


18G9. 


1884. 


75.675 

11.78 

1.12 

1.855 
trace 

trace 
trace 

4.75 

0.60 
1.72 
0.63 

1.87 


77.820 

12.11 

0.82 

2.420 

trace 

trace 
trace 

0.44 
0.37 

0.98 
1.04 


48.80 

39 00 
5.33 

1.22 

0.86 

1.28 


49.70 

34.00 

6.00 

1.75 




i—4 

i—4 

>-] 


8 
[5 
I 
.7 


34.66 
20.19 

1.21 

0.83 

0.96 


81.2 
12.15 

1.075 


-10.15 
-1.202 



All the crusts run much lower in silver than any from American 
refineries. Those from Altenau and Lautenthal retain more lead 
than any of the others, as they are to be melted down again in a 
kettle and decomposed by steam, which could not be satisfactorily 
done if the liquation had been carried any farther. 

It is usually stated that the total zinc required is added in three 
separate portions : first f, then \, and finally the remaining J^. 
While this corresponds approximately to actual practice, it cannot 
be implicitly followed, as the amount of the successive additions 
must depend on the assay of the bullion. The rate at which the 
silver in the lead decreases with the additions of zinc is influenced 
by a number of circumstances that cannot be determined in ad- 
vance. Each refinery has its own table to show che amount of 
zinc that shall be added if the bullion assays a certain number of 

1 Ztschrft. f. B. H. u. S. W. i. P., xxviii., p. 262. 

2 Ibid. 

3 Berg- und Buttenm. Ztg., 1875, p. 129. Engineering and Mining Journal, 
March 17, 1877. 

4 Ztschrft f. B.B.u. S. W. i. P., xxxiv., p. 92. 



328 DESILVERIZAT10X OF BASE BULLION. 

ounces of gold and silver to the ton. Some of these tables are very 
complicated. The following is one of the simplest. By it, the gold 
and silver from a 30-ton kettle are extracted separately with three 
zinc-additions. 

Table of Zint-additioxs for Gold. 
Up to 0.10 ounces gold per ton, 250 pounds of zinc. 



0.10 — 0.30 
0.30 — 0.50 
0.50 — 0.70 
0.70 — 0.90 



300 
350 
400 
450 



etc. etc. 

It is to be noted that gold and copper are extracted from the 
lead without saturating this with zinc. For instance, 30 tons of lead, 
taking up 0.6 per cent, of zinc, require 360 pounds of zinc, while 0.30 
ounces gold per ton are extracted by an addition of 300 pounds. 

When the gold-crust has been skimmed off the kettle, 500 
pounds of zinc are added for bullion assaying from 150 to 250 
ounces silver. With bullion running as high as 300 or 400 ounces, 
550 pounds are given. After removing the first silver-crust the 
kettle will assay from 10 to 50 ounces silver per ton, generally from 
30 to 40 ounces ; in exceptional cases it may run as high as 70 
ounces. The second and final addition of silver-zinc, varying from 
400 to 600 pounds, will'reduce the silver-contents of the kettle down 
to a trace, even if it has been as high as 70 ounces. Bullion run- 
ning low in gold and silver, say from 0.05 to 0.10 ounces gold and 
from 50 to 125 ounces silver to the ton, receives one gold-zinc and 
one silver-zinc, the resulting lead running less than 0.2 ounces 
silver to the ton and generally a trace. 

Latelv Edelmann and Rossler l have carried on an interesting 
series of experiments with the object of concentrating the silver in 
a richer alloy than usual. The ordinary silver-crust, they say, con- 
sists of a small quantity of zinc-silver alloy distributed through 
zinc-bearing lead which is partly oxidized. As the oxidized part 
obstructs a satisfactory separation of the zinc-bearing lead from 
the zinc-silver alloy, the improvements must consist in preventing 
the formation of oxides ; then the direct output of desilverized 
lead would be increased, an alloy rich in silver produced, and the 
consumption of zinc reduced. They suggest heating the lead, freed 
from copper by means of zinc, up to incipient redness (500° C), 

1 Berg- und Huttenm. Ztg., 1890, pp. 245, 429; 1891, p. 123. Engineering 
and Mining Journal, November 15, 1890; April 4, May 16, 1891. 



PARKED PROCESS. 329 

introducing the zinc, heated to a higher temperature than the lead, 
at the bottom of the kettle, which does away with stirring and 
skimming the little oxidized crusts from the surface, keeping the 
first rich crusts separate from the following ones running lower in 
silver, and utilizing these with fresh zinc. The addition of alu- 
minum to the zinc, in some of their experiments, had a favorable 
effect in preventing oxidation, and thus assisted in obtaining crusts 
very rich in silver. At a Belgian refinery ' crusts running 20 per 
cent, silver have been obtained by adding aluminum with the zinc, 
and experiments of electrolyzing these rich crusts are under way. 
In this connection Rosing's electrolytic treatment of silver-crusts 2 
is of special interest. Other suggestions regarding improvements 
in desilverization have been made by Rosing, 3 Honold, 4 and 
Schlapp.' 

§ 101. Desilverizing-Kettles.— The kettles (Figure 231) used 
for desilverization are spherical in form. They are from 3 feet to 
3 feet 4 inches deep, and their diameter varies according to the 
required capacity. Most kettles used to-day hold 30 tons; the 
dimensions of such a kettle are shown in the figure. A circular 
kettle of a greater capacity would present difficulties in working. 
In the few instances where kettles holding over 30 tons are used, 
they are made oblong and have rounded ends. Such a kettle, e.g., 
holding 55 tons of lead when actually filled, and 45 tons when filled 
to 5 inches from the top, is 12 feet long, V feet wide, and 3 feet 3 
inches deep. 

The kettles are all of cast iron. The iron used should be dense 
and strong, but not hard. Lake Superior No. 2 iron, which is 
neutral, strong, and tough, mixed with car-wheel iron, giving the 
necessary density, furnishes a material filling every requirement. 
Kettles are best cast bottom down, as it gives greater density. 
The other way is easier for the foundry. In any case they should 
be cast in brick-work if a smooth surface is to be obtained, which 
is necessary for the scraping of the kettle ; any blisters, besides 
weakening the kettle, are liable to assist the corroding effect of 

1 Private communication, December, 1891. 

2 Berg- und Euttenm. Ztg., 1886, pp. 463, 478, 488; abstract in Engineering 
and Mining Journal, May 22, December 25, 1886. 

3 Chemiker Zeitung, 1889, p. 1059. 

4 Berg- und Euttenm. Ztg., 1890, p. 187; 1891, p. 342. Stahl und Eisen, 1891, 
p. 152. 

6 American patent, No. 380,524, April 3, 1888. 



330 DESrLVERlZATION OF BASE BULLION. 

the zinc, and to retain particles of zinc-crust which may enrich 
already desilverized lead. Kettles have often been mado 2 J inches 
thick at the bottom, tapering to l-J or 2 inches at the rim. At 
present they are made of uniform thickness throughout, and rarely 
over 1J inches in thickness. Such a kettle lasts from one to one 
and a half years, being in continual use. Steel kettles are not 
found in American desilverizing works. They have come into use 
in European works, where the lead is refined by means of steam in 
the same kettle in which it was desilverized. A kettle is usually 
suspended by its rim, which rests on a circular cast-iron ring cover- 
ing the top and sides of a brick wall. Figures 225, 229, 230 show 
three supporting rings ; see Figure 225 for the desilverizing kettle, 
Figure 229 for the liquating-kettle, and Figure 230 for the 
liquated-lead kettle. The casting (Figure 225) consists of four 
separate pieces, shown in section by Figure 226. They are 
fastened together by bolts passing through flanges, as seen in 
Figure 227. The casting rests on the working-platform of the 
kettle, as drawn in the front elevation (Figure 205). It is made 9f 
inches wide, and encloses the 9-inch wall, which rises 18 inches 
above the main brick-work. With many kettles the support of the 
rim consists of a circular iron ring, covering only the top of the 
brick-work. In such cases this must be thicker than 9 inches, if it 
is not to give way to the pressure of the weight of the kettle filled 
with lead. The side- wall thus often reaches a thickness of 18 
inches, the iron support-ring either entirely covering it or leaving 
2 or 3 inches exposed. Miiller ' gives it as his experience that a 
kettle lasts longer if suspended from a rib cast on the kettle at half 
its depth. This rib would then form the partition- wall between 
the fire-place and the encircling flue, which simplifies the construc- 
tion of the brick-work. The place from which the kettles are fired 
is seen in the front elevation (Figure 205). The horizontal section 
(Figure 208) shows the plan of the brick-work with the ash-pits of 
the three kettles. 

Figures 206, 207, 209 give more detail of the brick-work, and 
show the road of the products of combustion from the grate to t?he 
flue leading to the chimney. In the desilverizing-kettle the flame 
goes from the fire-place d (Figures 206, 209), first back and up- 
ward ; it then passes around the kettle to the right, in a circular 
flue (as indicated by the arrows), and leaves this at e, entering a 
vertical flue leading to the main chimney. In the liquating-kettle the 
1 Berg- und Huttemn. Zty., 1889, p. 218. 



■ 






v 



.33MAR-'- 









PL 









i 

















j hot-o: 




1 3JT1 







; 

3HT 



^1 ,C 












ai 







PARKES' PROCESS. 331 

products of combustion go from the fire-place /(Figures 207, 209), 
after passing under the kettle, straight into the flue g. The gases 
from beneath the liquated-lead kettle h (Figure 209) go to the 
left, and join those of the liquating-kettle. 

Desilverizing-kettles were formerly emptied by a discharge- 
pipe cast in the bottom of the kettle and running out through 
the brick-work. It was closed either by a slide-valve on the out- 
side (similar to the lead discharge of the Luce-Rozan crystallizer, 
Figures 191-193), or by a clamp and thumb-screw on the inside. 
At present the Steitz siphon is in common use, and is preferable 
to any other means for emptying a kettle. A common form of it 
is shown in Figure 235. It consists of a piece of gas-pipe a from 
2 to 2J inches in diameter, bent so as to reach from the rim of the 
kettle to the bottom. Here it has an elbow b screwed to it to 
prevent the lead column from breaking. To the other is attached, 
also by an elbow c, the vertical section-arm d, having a cast-iron 
stop-cock e near the lower end. The siphon discharges the lead 
into a cast-iron trough of f-inch iron, which carries it into the 
refining-'furnace. 

§ 102. Liquating Apparatus. — In connection with the desilver- 
izing-kettle must be discussed the apparatus required for liquating 
the zinc-crusts. Two sorts are in use. The first is shown in Fig- 
ures 232 and 233. It consists of a shallow kettle placed on the 
same level with the desilverizing-kettle and close to it. The bot- 
tom is convex in the centre, in order that the discharging-spout 
may be shorter than would be possible if the kettle had the usual 
spherical form. Formerly a perforated cast-iron disk (an old 
skimmer) was placed inside the kettle, over the opening into the 
spout, to prevent particles of crust from passing off with the liquid 
lead ; this has been given up at some works, as the perforations 
easily become clogged, and when open do not prevent fine particles 
from being carried off by the liquated lead. This runs from the 
spout into a small spherical kettle, whence it is bailed out after it 
has been skimmed. 

The drawings given show for every desilverizing-kettle one 
liquating-kettle with its liquated-lead kettle. This presupposes 
that no distinction is made between gold-crusts and silver-crusts, 
with the result that all the silver produced contains gold, and has to 
be parted. Where the crusts are kept separate the desilverizing- 
kettle will have a liquating-kettle on either side, one for the gold- 
crust, the other for the silver-crust, the liquated crusts as well as 



332 DESILVERIZATION OF BASE BULLION. 

the liquated lead from the two small spherical kettles being kept 
separate. Only the silver resulting from the gold-crust will then 
have to be parted. With a plant where the two crusts are kept 
separate, the distance between the centres of the two desilverizing- 
kettles of 32 feet, as shown in the general plan (Figure 196), will 
be too small ; it will have to be enlarged sufficiently to leave a 
passage-way between the two small liquated-lead kettles. 

The second apparatus is a reverberatory-furnace placed on the 
floor of the desilverizing-kettles. It may be built as follows. The 
hearth is a cast-iron plate, 10 by 5 feet, slightly trough-shaped, 
and having a rim 4 inches high along the sides. To the lower end 
a spout is attached, which ■ discharges the liquated lead outside of 
the furnace into a small liquated-lead kettle of the same form as 
the one described, the rim of which is on a level with the working- 
floor. The plate has an inclination of 3 inches. It lies on a bed of 
brasque, which is tamped into a wrought-iron pan supported by trans- 
verse rails resting on two longitudinal walls. The furnace has two 
working doors on one side. The writer prefers the reverberatory- 
furnace to the liquating-kettle, as he obtained from it in a shorter 
time than from the kettle a dryer crust, and thus a richer bullion, 
without driving some of the silver from the crust back into the 
liquated lead, as often happens with the kettle. The reason for- 
the better result in the reverberatory furnace is probably that the 
lead can be gradually eliminated at a slowly increasing tempera- 
ture in a reducing atmosphere, which prevents the oxidation that 
always takes place in a liquating-kettle, even if it be covered with a 
sheet-iron plate, on raising the temperature to the required degree. 
Supposing, however, the liquation to be equally good with the two 
apparatus, there remains this advantage for the reverberatory, that 
the liquation of the silver-crusts of several systems can be performed 
continuously in one furnace, w^hich, being separated from the desil- 
verizing-kettles, does not disturb the work there, and which collects 
all the rich crusts into one place, whence they are delivered through 
a chute into the bins of the retort-room. With a small plant a 
liquating reverberatory furnace of any reasonable dimensions would 
not have sufficient work to do to pay for the extra labor. Here the 
liquating-kettle must be used, even if it obstructs to some extent 
the work in the desilverizing-kettle. The preference for the rever- 
beratory furnace is not general, as several important refining works 
adhere to the liquating-kettle. 

The practice of liquating zinc-crusts in a spherical kettle with- 



PARKES' PROCESS, 333 

out a lead-discharge and removing the liquated crust floating on the 
liquid lead with a skimmer is antiquated, if the crust is to be dis- 
tilled, and justly so, as it is impossible to obtain in this way a dry 
crust that is satisfactory. If the temperature of the lead be raised 
sufficiently high to obtain a dry crust, a considerable quantity of 
it will be redissolved by the lead. It will rise again to the surface 
when the lead cools, but it will be rich in lead, and at the same time 
low in silver. It must, therefore, be returned to the next charge 
that is to be liquated, instead of going directly to the retort, as it 
would with either of the two apparatus just discussed. 

§ 103. Mode of Conducting the Desilverization. — This is as 
follows. To the lead in the kettle, which represents, after the kettle- 
dross has been removed, about 90 per cent, of the bullion charged 
into the softening-furnace, is added lead obtained by liquating the 
gold-crust from a previous charge. It is then heated above the melt- 
ing-point of zinc (412° C.) and receives the first zinc to remove the 
gold and the copper. The heating takes a very short time ; in half 
an hour from the time when the kettle is drossed the zinc will 
usually be melted down. 

It is customary to place the slabs of zinc on top of the lead and 
to heat this until the zinc has become thoroughly melted, when the 
stirring-in can begin. Objections have been and are raised to this 
method of adding the zinc, which, floating on the lead, is exposed 
for some time to the oxidizing action of the air. To avoid this, 
the zinc has been enclosed in a perforated iron box, which is forced 
down to the bottom of the kettle and held there by an upright iron 
rod fitting into the centre of a piece of flat iron which reaches 
across the kettle, and is fastened to it by set-screws. There can be 
no question that the zinc will melt more quickly at the bottom of 
the kettle, will not be exposed to the air, and rising upward in thin 
little streams will combine more readily with the lead and the 
silver, and will require less stirring-in than if melted while floating 
on the lead. French refineries use this method frequently. Accord- 
ing to Edelmann and Rossler, 1 it is advantageous to melt the zinc 
in a separate vessel and to pour it into the lead, as it is then taken 
up quickly by the lead. It seems, however, doubtful whether less 
dross is formed than if the zinc is melted down floating on the lead. 
Figures comparing the results of these three methods of incorporat- 
ing the zinc would be of interest. 

In pr^.er to bring the melted zinc into intimate contact with the 

1 Loc.cit 



334 DESILVERIZATION OF BASE BULLION. 

argentiferous lead it is stirred in for from one-half to three-quarters 
of an hour. When this is done by hand, a paddle is used, consisting 
of a perforated disk 12 inches in diameter, riveted to a handle 6 
feet long, having a cross-piece as hand-hold. Two men standing 
opposite each other do the stirring. They insert the paddles verti- 
cally at the rim of the kettle, push them downward towards the 
centre, raise them, using the rim of the kettle as a fulcrum, and draw 
them, with the disks gliding on the lead, from the centre towards 
the periphery, giving the lead a rotary motion which they reverse 
every five minutes, thus insuring an intimate mixing of zinc and 
lead. This stirring-in has always been hard for the workmen, and 
various mechanical stirrers have been devised, but stirring-in by 
hand remained the universal method until steam-stirring came into 
use. The effect of steam when introduced into lead containing 
zinc varies greatly according to the temperature of the lead. 

1. If the temperature of the lead be below the melting-point of 
zinc, i.e., the temperature when the kettle is skimmed, the steam 
will bring to the surface a zinc-crust, and with it some of the silver 
contained in the lead. 

2. If the temperature be slightly above the melting-point of 
zinc (stirring-in time), the steam will cause a thorough mixing of 
zinc and lead. 

3. If the temperature be between a dark-red and an incipient 
cherry-red, the steam will cause a scum to rise, containing about 
3 per cent, of zinc, which does not, however, take any silver away 
from the lead. 1 

4. If it be a clear cherry-red, the zinc will decompose the 
steam ; the resulting zinc oxide (mixed with lead oxide) collects as 
a powder on the surface of the lead. 

The steam must be absolutely dry if violent explosions are to 
be avoided. The condensed water is separated by a steam-trap 
placed beneath the working-platform. To the vertical pipe coming 
through the platform is fastened by means of a coupling, a small 
piece of pipe, to which is joined by two elbows (with a nipple in- 
tervening) the horizontal pipe which reaches to the centre of the 
kettle. An elbow connects it with the vertical pipe, that will 
reach 24 inches into the kettle when in place. Before the steam- 
valve is opened, the vertical pipe is turned up, in order that the 
steam may first pass out into the air to warm the pipe and to insure 
the expulsion of any condensed water. The pipe is then turned 
1 Rosing, Z. f. B. K u. S. W. i. P., xxxvii., pp. 76 and 77. 



PARKES' PROCESS. 335 

down and weighted with a bar of lead to keep it in place. When 
the steam is turned on, the waves of lead caused by the first ascend- 
ing bubbles will drive the zinc floating on the lead toward the rim 
of the kettle, and hardly any of it would become incorporated with 
the lead if it were not pushed towards the centre of the kettle to 
be drawn into the lead by the downward current close to the steam- 
pipe. The zinc-crusts that rise to the surface soon after stirring-in 
has begun are also pushed toward the centre, that they may take 
up more silver. Thus the zinc and then the crusts pass down at 
the centre, and come up again nearer the circumference of the 
kettle, whence they are again pushed towards the centre. 

The tool used for this purpose is a wooden hoe, consisting of an 
inch board, 12 by 18 inches, into the centre of which is inserted 
an inch lath from 8 to 10 feet long. 

Towards the end of the stirring-in, which lasts from one-half 
to three-quarters of an hour, the fire beneath the kettle is removed 
or damped with wet slack coal, the fire-doors and ash-pit doors 
and the damper in the flue are throwm wide open, and the kettle 
is allowed to cool. At St. Louis, A. Meyer 1 tried to hasten the 
cooling by means of water-cooled pipes bent to the shape of the 
kettle, but his apparatus has found no favor, as the pipes did not 
cool the lead enough to pay for all the trouble and inconvenience 
connected with their use. At St. Louis they were given up many 
years ago. A better means of hastening the cooling, although 
not much used, as it injures the brick, is to sprinkle water from a 
hose against the inside walls. If the zinc is stirred in by steam, 
its use will be found effective afterward in cooling. 

After from two to three hours the lead has cooled down so far 
that the crust begins to adhere to the sides of the kettle. It is 
then removed with a skimmer, the disk of which, made of J-inch 
iron, is from 14 to 18 inches in diameter; and the handle, of 1^-inch 
iron gaspipe of the same weight, is 7 feet long, having a cross-piece 
as hand-hold. The work is facilitated by suspending the skim- 
mer by a hook, which acts as a lever. Two men work together. 
One man pushes with a wooden hoe the crust towards his partner, 
who takes it up with the skimmer. Before discharging it into the 
liquating kettle or the mould, it is important that the skimmer be 
well jerked several times, in order that as much lead as possible 
may be drained off into the kettle and the crust obtained become 

1 Mining and Scientific Press, 1882, vol. xliv., No. 5 ; Berg- und Huttemn. 
Ztg. } 1882, p. 391. 



336 DESILVERIZATION OF BASE BULLION. 

dry. Toward the end of the operation both men have to work 
very slowly and carefully to avoid pushing the crusts back again 
into the lead, which would retard the work very much. When the 
crusts have been removed from the surface, the alloy adhering to 
the sides of the kettle has to be brought to the surface, which is 
done by scraping them first with a chisel-pointed bar and then 
with a wooden lath. The blade of the bar is of steel, 2 by 4 
inches, and the handle of 1-inch round iron. Skimming and scrap- 
ing are generally repeated twice, after which no more crusts will 
rise to the surface. It takes about an hour to perform this opera- 
tion. When finished, the fire under the kettle is again started, and 
this heated to melt down and stir in the next zinc. The time 
required for heating up varies from one to two hours, according to 
the amount of lead (obtained from liquating zinc-crusts) that is 
added to the kettle. 

A sample is taken from the lead to see whether all the gold has 
been extracted and how much silver. It is well to cupel eight 
samples of an assay-ton each, and dissolve the silver buttons 
together to ascertain if all the gold has been taken out. 

The gold-crust obtained by the first skimming is collected either 
in one of the liquating kettles or, if the reverberatory furnace is 
in use, in flat, slightly conical moulds that have been whitewashed. 

While the kettle is being heated up, liquated lead from the 
first silver-crust of the previous charge is added, and when the lead 
is sufficiently hot, the zinc for the extraction of the bulk of the 
silver. The poor second silver-crust obtained from the previous 
charge is used with third crust, if there was any, and fresh zinc. 
The operations are the same as for the gold-crust. 

The second and third crusts are also sometimes collected in 
moulds, but oftener the skimmer is discharged on a whitewashed 
iron plate. There the crusts remain until needed, when one end 
of the plate is raised by block and tackle, and the two crusts are 
slid into the kettle together. Some works discharge only the 
second silver-crust on the plate, while the third goes into moulds. 

After each skimming, samples are taken, and % assay-ton is 
assayed for silver to check the progress of the desilverization. 
After the last silver-crust has been removed, the assay should show 
0.2 ounces silver per ton, or less if corroding lead is being made. 
Should it prove to be slightly higher, say 0.4 or 0.6 ounces, the 
introduction of steam at the low temperature of the kettle after 
the last crust has been taken off will be effective in causing more 



PARKES' PROCESS. 337 

silver-bearing zinc-crust to be given off by the lead, as stated above. 
If the steam is used from one-half to three-quarters of an hour, 
the silver-contents of the kettle will be reduced, and thereby gener- 
ally an entire zincing saved. 

In a 30-ton kettle five hours are allowed for each zincing if the 
silver is extracted by four additions of zinc. The first five hours 
include the time the softened bullion is being run into the kettle, 
the last five hours the time during which the desilverized lead is 
being syphoned into the refining-furnace. When the lead comes 
from the softening-furnace it is usually hot enough to melt the first 
zinc quickly, and the melting down of the last zinc and skimming 
the final crust require but little time, so that the hour necessary to 
tap the softening-furnace and empty the kettle can be included in 
the standard five hours. With a skilled crew it is possible to 
desilverize a 30-ton kettle with four zincings in eighteen hours, but 
every moment has to be utilized to accomplish it. Some works, 
therefore, aim to extract the precious metals with three additions 
of zinc, and give six hours for each addition, keeping the gold and 
silver crust separate. When gold and silver are extracted together, 
a 45-ton kettle can be desilverized with three additions of zinc in 
the same length of time. The reduction of time may be still 
further increased by the substitution of oil or gas for coal as fuel. 

The weight of the single crusts varies considerably, and also the 
amount of silver which they take up. As an illustration, 1,500 
tons of softened bullion, containing 170 ounces silver and 0.5 
ounces gold to the ton, gave 5 per cent, of liquated gold-crust, 
required 0.3 per cent, zinc, and reduced the silver-contents 15 
ounces silver per ton ; the liquated silver-crust was 8 per cent., 
required 1.5 per cent, zinc, and reduced the silver-contents to 30 
ounces per ton. The 1.5 per cent, represents the entire zinc added 
to extract the silver. The second silver-crust reduces the assay of 
the kettle to about 3.0 ounces per ton, and the third to 0.2 ounces 
or less. Where only three zincs are added, the silver-contents are 
reduced by the second silver-zinc from 30 ounces silver to the ton 
to 0. 2 or less ounces. From the zinc added to the kettle part is 
recovered by distilling the zinc-crusts and makes the amount 
actually consumed lower. This was one per cent, for a year's run 
for softening bullion in the kettle, which averaged 150 ounces 
silver and 0.5 ounces gold to the ton, with the usual amount of 
copper. The coal required for desilverization, including liquation, 
is about 54 pounds for a ton of unsoftened base bullion. 



338 DESILVEBIZATION OF BASE BULLION. 

In connection with desilverization may be mentioned an excel- 
lent custom ' of breaking up unwieldy old kettles that have been 
set aside on account of leakage or corrpsion. This is done by till- 
ing the kettle with water and suspending a dynamite cartridge 
from a floating board so that it nearly touches the bottom. When 
this is exploded the water will be thrown up into the air and the 
kettle broken into five or more pieces, which can be easily handled 
and shipped to the foundry. 

§ 104. Method of Liquating. — The liquation of the gold and 
silver crusts takes place, as stated in § 102, either in a liquating- 
kettle or a reverberatory furnace. It is essential that the temper- 
ature be raised gradually and that small quantities only be 
liquated at one time. If in the liquating-kettle the temperature is 
raised quickly, some of the zinc-crust will be taken up by the 
melted lead and carried off into the small liquated-lead kettle, 
whence it is again skimmed off as a crust rich in lead, to be treated 
with the next batch of crust. The kettle must therefore be large 
enough, as shown in Figures 232 and 233, to hold the silver-crust 
in not too thick a layer. With the reverberatory-furnace a charge 
is introduced through the door near the flue, and gradually moved 
toward the fire-bridge, whence it is drawn out through the second 
door upon an iron plate let into the floor. By moving the crust 
from the coolest part of the furnace to the hottest, a gradual 
elimination of the lead takes place, and there is little danger of 
dissolving the crust, which grows less and less fusible, in the lead, 
which runs off as soon as melted. 

The zinc-crust, when still hot, can be easily broken up into 
small pieces. When removed from the kettle or the furnace it is 
spread on an iron plate and reduced to nut size, and smaller, by 
flattening with the back of a shovel and working with a rake. In 
this form it is readily charged through the narrow neck of the 
retort. Another way to break up the liquated crust is to transfer 
it from the liquating apparatus into a cast-iron box (24 by 18 
inches, with a rim 3 inches high), the bottom of which has a series 
of slits (18 by 2 inches), alternating with ribs of the same width, 
the casting being \ inch thick. With a liquating kettle this box 
requires a frame as support ; with the reverberatory furnace it is 
attached to the door frame. 

The lead recovered by liquating is from 40 to 60 per cent, of 
the weight of the crust charged. In desilverizing 250-ounce bullion, 
1 Lautenthal, Smelting and Refining Works. Private notes, 1890. 



PARKES' PROCESS. 339 

the lead from the gold-crust assays from 100 to 200 ounces silver per 
ton, that of the first silver-crust from 30 to 40 ounces. The lead li- 
quated from the gold-crust is added to the desilverizing-kettle before 
the gold-zinc is given, that from the silver-crust before the first 
silver-zinc. 

§ 105. Refining Desilverized Lead. 

Introductory Remarks. — The desilverized lead retains, after 
the last crust has been removed, from 0.6 to 0.7 per cent, zinc, 
according to the temperature that prevailed at the last skimming. 
To remove this, and also small quantities of arsenic and antimony 
that were either not entirely taken away during the softening, or 
that were introduced with the zinc used in desilverizing, the lead 
must undergo a refining process. 

From the desilverizing-kettle the lead is siphoned off into the 
apparatus used for refining, which is in most American refining 
works a reverberatory furnace, in a few instances only a spherical 
kettle. 

The siphon (§ 101) is heated and filled by immersing it in the 
kettle, the stop-«ock being open. When it has attained the tem- 
perature of the lead and is filled entirely with it, the stop-cock is 
closed with a key, the longer arm taken out and suspended, and the 
shorter one held down in the lead. The stop-cock is then opened, 
and the lead runs out into a cast-iron trough, which discharges 
into the refining-apparatus. To keep the siphon in place it is 
weighted by a couple of bars of lead. Should the lead-column 
break when the kettle has, for example, been half emptied, and it 
not be possible then to fill the siphon again in the usual way, it 
will be necessary to invert and fill it by ladling; for this purpose 
an iron funnel may be used to avoid delay. But the breaking of 
the lead-column is a very rare accident. 

§ 106. Refining in the Reverberatory Furnace. — The gen- 
eral construction of the reverberatory furnace used for refining is 
the same as that for softening. Formerly it was customary to 
make the refining-furnace smaller than the softening-furnace, in 
order that it might correspond to the smaller amount of lead it has 
to treat. For instance, the hearth of a softening-furnace of a 25- 
ton kettle was made 13 by 9 feet and 13 inches deep, having 13- 
inch side-walls; that of the refining-furnace 12 by 8 feet, and of the 
same depth, with the same thickness of walls. At present it is 
commoner to give the refining-furnace the same dimensions as the 
softening-furnace, only making the hearth slightly shallower, thus 



340 DESILVERIZATION OF BASE BULLION. 

simplifying the iron parts. If in the softening furnace, shown in 
Figures 198-204, the depth of the hearth is reduced 3 inches, it 
will have the capacity necessary to receive the desilverized lead 
from the kettle, which will fill it to just below the skimming doors. 
In some refining-furnaces the lowest point of the hearth is placed 
beneath the flue, as in the softening-furnace described; in others, 
below the central skimming door. By referring to the general 
plan (Figure 196), it will be seen that the lead is discharged from 
below the door next to the flue. The arrangement for tapping is 
usually the same in both furnaces. If, however, the refined lead 
is to be conveyed directly from the furnace into the moulds, as is 
still found in a few instances, instead of being moulded from the 
" Merchant-kettle," the tap will be slightly different from that of 
the softening-furnace (see § 108). 

The mode of operating is similar to that in the softening-fur- 
nace. When the furnace is filled, the fire is urged, as a high heat 
is required to burn off the zinc. This is partly volatilized and 
carried off with the fumes, and partly oxidized and scorified by 
the litharge which forms at the same time. Some refiners add 
lime to the charge, as in softenings After heating; about four hours 
the surface of the lead will be covered with a heavy litharge-like 
skimming. The doors are thrown open, the skimming is removed, 
and the second heat given, after which cooling and skimming are 
repeated. A third heat is often necessary to slag the last traces 
of zinc and antimony. When these are completely removed, the 
litharge drawn from the surface of the lead by means of a rabble 
should be in large, thin plates. It should, while hot, have a bright- 
yellow color when seen in bulk, and a greenish-yellow when held up 
to the light, but not one brown spot (antimony) should be visible. 
If these large flakes of litharge should become dark or show spots 
after having been exposed for some time to the air, the lead is not 
sufficiently refined to satisfy the requirements for corroding-lead. 
In firing the refining-furnace it is essential for the temperature to 
be raised quickly to the necessary intensity, and kept there. If it 
is allow r ed to fall even a little, the burning-off of the zinc will be 
greatly retarded, and with it the dependent operations. 

As in softening the base bullion, steam was introduced through 
pipes to expose more surface to the oxidizing action of the air, so 
in refining, steam is also used, though not so often now as formerly. 
In addition to the mere mechanical effect of stirring, it acts 
chemically by being decomposed by the zinc (§ 103), and thus 



PARKED PROCESS. 341 

hastens a great deal the elimination of zinc from the lead ; but, as 
in softening, the side-walls of the furnace are more easily attacked, 
and the percentage of skimmings increased. 

The time required for refining with a 30-ton plant is about 
fourteen hours ; the refining skimmings amount to from 4 to 5 per 
cent, of the bullion charged, and contain about 90 per cent, of 
lead. The coal consumed in refining is about 120 pounds per ton 
of unsoftened bullion. 

§ 107. Refining in the Kettle. — The second method of refining 
desilverized lead is the one invented by Cordurie, who introduced 
steam into the lead heated up to cherry-redness, the oxidized zinc 
collecting in the form of a powder on the lead. As the air cannot 
be excluded from the surface, and as it is also carried in by the 
steam, some of the lead is oxidized, and the pulverulent yellow 
mass floating on the surface consists of a mixture of lead and zinc 
oxide and finely divided shots of lead. The composition of these 
oxides when taken from the kettles at the Lautenthal 1 Smelting 
and Refining Works (Prussia) is 



Sb 2 5 , 


1.893 


Fe 2 O s , 


0.986 


ZnO, 


23.775 


PbO, 


37.933 


Pb, 


34.236 



The base bullion at Lautenthal is desilverized without previous 
softening ; hence the high percentage of antimony. 

The larger shots of lead of the oxides are separated by screen- 
ing, the finer ones by washing over an inclined plane. The impal- 
pable powder, forming 15.44 per cent of the whole, floats off and is 
settled in vats, dried in a reverberatory furnace, and forms a 
reddish-yellow paint of good covering power. It consists of 

67-60 per cent. ZnO, 
33-40 per cent. PbO. 

The residual shots of lead are smelted at intervals for a second- 
class lead, as they contain some antimony. 

At Lautenthal 2 the cast-iron kettles (5 feet 41 inches in 
diameter, and l 2 feet 9 J inches deep, holding 12-^ tons of lead) 
are heated, after the desilverization is finished, in four hours, to a 

1 Z. f. B. H. u. S. W. i. P., xxxviii. p. 272. 

2 Private notes, 1890. 



342 DES1LVER1ZATI0N OF BASE BULLION, 

cherry-red ; superheated steam having a pressure of from 29 to 36 
pounds per square inch is then introduced through a cast-iron pipe 
bent to the form of the kettle, so that the steam enters at the 
bottom. After two hours all the zinc has been oxidized. In order 
to decrease the loss of heat by radiation, to keep off the air, and to 
prevent the oxides from being lost, the kettle is covered by a 
movable sheet-iron cylinder, 7 feet 8 inches high, which has near 
the lower rim two opposite doors (4| inches square) and the 
opening for the steam-pipe. It ends in a conical hood which leads 
the vapor and dust through a sheet-iron pipe (1 foot If inches in 
diameter) into the main flue, terminating in a dust-chamber. The 
cylinder with its hood and pipe is suspended by a running differ- 
ential pulley. 

For every 100 pounds of unsoftened bullion 4.67 pounds of 
paint are produced, which is higher than the percentage of skim- 
mings in the reverberatory furnace, as the paint forms but a small 
part of the total lead taken out of the kettle. 

The great drawback of this method is the wear and tear in 
kettles. According to Schmieder, 1 at Tarnowitz, where the lead is 
very free from antimony, cast-iron kettles hold out only for twenty 
charges, while steel kettles are good for ninety charges. The life 
of cast-iron refining kettles varies greatly ; extreme figures are 30 
days and 120 days. Of kettles from the same foundry, cast under 
apparently the same conditions, one will last only a short time, 
while another, in the writer's experience, has lasted over a year. 

As by the use of steam a considerable amount of air is carried 
into the lead and some of this oxidized, Rossler 2 tried to replace it 
by different gases. With carbonic acid, the lead being heated to 
700° C, in a short time all the zinc was converted to white oxide, 
and could be skimmed off from the lead. As carbonic acid cannot 
be easily obtained pure to be used for such a purpose, he tried a 
mixture of carbonic acid and nitrogen, obtained by pressing air 
through a cylinder filled with glowing coal. The result was a gray 
powder, in which some of the zinc was present as metal in a finely- 
divided state. By the use of a mixture of carbonic oxide and 
nitrogen, drawn from a gas-producer, a powder of a darker gray 
was obtained, containing up to 75 per cent, of its zinc as metal in 
a finely-divided state, the rest being oxidized by the carbonic acid 
present. The refined lead was entirely free from zinc if the tern- 

1 Berg- und. Huttenm. Ztg., 1887, p. 377. 

2 Berg- and Huttenm. Ztg., 1890, p. 248. 



PARKES 1 PROCESS. 



343 



perature was kept above 700° C, otherwise a zinc-crust formed on 
top of the lead. 

§ 108. Moulding from the Refining Furnace. (Figure 236.)— 
The moulding of lead was formerly universally done by ladling 
it from a kettle. This has, however, become antiquated. In 
some works the lead is moulded directly from the reverberatory 
furnace in which it has been refined. At the lowest part of the fur- 
nace, on the side away from the desilverizing-kettle, a piece of 2-inch 
gas-pipe is screwed into the pan if of cast-iron ; if of wrought- 
iron, into the flanges placed on either side and fastened together 
with bolts. If the furnace has a wrought-iron jacket, the pipe b 
(Figure 236) is screwed into the cast-iron plate placed between the 




ITig. 235. 



10 11 12 



APPARATUS FOR THE MOULDING OF LEAD. 
i( 2 Fig. 236. 




two sides of the jacket at the tap-hole. The pipe b is joined by a 
cast-iron stop-cock c to a T, whose horizontal arm is closed with 
a plug d. To the vertical end of the T is attached a nipple e, 
with an elbow / at the lower end, into which is screwed a long 
pipe g (from 7 to 10 feet), which can be moved horizontally by the 
arm h, while it discharges the lead into moulds placed in a semi- 



3U 



DESILYERIZATIOX OF BASE BULLIOX 



circle, the centre of which lies beneath the nipple. The moulds 
commonly used now (Figures 237-239), differ from the ordinary 



LEAD MOULD ON WHEELS. 

Fig. 237. 



PLAN. 



9**- 



12 9 6 3 

I i i I i i I i i I i , I 



Fig. 338. 

SECTION ON THE LINE A B. 



' 



Fig. S39. 

FRONT VIEW. 






blast-furnace moulds in that one end rests on two wheels a, while 
the other has a leg b. The lip of the mould above this has a 
hole c. By passing a hook through it and tilting the mould, it 
is run away, the bar tipped out, and the mould then quickly 
returned to its former place. At some works the moulds are 
made large enough to hold three bars of lead. The lip then has, 
instead of the hole c, a rectangular socket running horizontally, into 
which is inserted a slightly bent iron handle to move and tilt the 
mould. 

When the furnace is to be emptied, a charcoal fire is started 
under the stop-cock, and the horizontal pipe immersed in the 
lead of the refining furnace to be warmed. It is then screwed 
into the elbow, the stop-cock is opened, and the lead run into the 
first mould of the semi-circle. This warming of the pipe is, how- 
ever, not necessary. If the stop-cock be opened entirely, the first 
lead arriving at the end of the horizontal pipe will still be liquid. 
After that, the cock will have to be slightly closed, as the moulds 
would otherwise fill too quickly for the man, who has also to attend 
to the skimming of the surface of the bars. This he does with 
two thin pieces of board the width of the mould, collecting the 
dross between them and dropping it on the floor. Another method 
is to rake off the dross on to the floor with a bent piece of hoop-iron. 
The former method gives a cleaner bar. 

One mould after another is thus filled. When the lead in the 



PARKES' PROCESS. 345 

first three or four moulds has solidified, it is chilled with water, 
the pigs are trimmed with a sharp, chisel-pointed bar, and the 
moulds run off to the wall of the lead-pit (Figure 196), where the 
lead is to be dumped before weighing, and brought back empty to 
their places. In this way from forty to sixty moulds that form 
the semi-circle are filled one after the other. 

This method of moulding has the advantage that the vertical 
distance between the shipping level of refined lead and the receiv- 
ing level of base bullion can be less than with the methods to 
be discussed ; then the filling of the moulds need not be con- 
tinuous, as it must with the siphon. With a kettle holding thirty 
tons of lead, the moulding can thus be given in contract to the four 
men in addition to the charging, softening, desilverizing, and 
refining. They do it in two separate operations, moulding part of 
the lead between the third and fourth zincings, and the rest while 
the last crust is rising, the total time required for moulding being 
six hours. The method has the disadvantage that the furnace has 
to be cooled down considerably for the lead to attain the right 
temperature for moulding. Thus after every moulding, two hours 
or more are required to heat up the furnace again for the next 
charge. This cooling and heating-up of the furnace with every 
charge cannot be good for the lining, therefore the moulding of 
the lead directly from the refining furnace has not found so much 
favor as might be expected. 

§ 109. Moulding from the Merchant-Kettle.— It has become 
more common to tap the refined lead from the reverberatory fur- 
nace into a kettle — the merchant-kettle — (Figure 196) heated from 
below, and to let it cool there till it has attained the correct tem- 
perature for moulding. In this case the tapping of the furnace is 
done in the same way as from the softening furnace. 

If the desilverized lead is refined by means of steam in a kettle, 
the moulding is done either from the refining-kettle or the lead is 
siphoned off into a merchant-kettle below, to store the lead until it 
is time for moulding, and thus have the refining-kettle ready for 
another charge of desilverized lead. 

Some refining works pole the lead in the merchant-kettle at a 
low temperature, under the impression that an especially fine grade 
of corroding-lead is thus obtained. Poling is, however, not neces- 
sary, as all the impurities not only can be but ought to be com- 
pletely removed in one operation, be it in the reverberatory furnace 
or in a refining-kettle. 



340 DESILVERIZATION OF BASE BULLION. 

As this poling at a low temperature is the best method of purify- 
ing the otherwise pure leads obtained from smelting the clean, 
non-argentiferous ores of the Mississippi Valley, a few remarks 
are in place. The gases and vapors from the wood stir up the lead 
and expose continually new surfaces to the oxidizing action of the 
air. Thus the small amounts of arsenic, antimony, copper, zinc, 
and iron are slowly oxidized and collect on the surface as a dross. 
A crutch serves to keep the stick of wood horizontally depressed 
in the molten lead. It consists of a piece of flat iron long enough 
to reach about 1£ feet over the rim of the kettle, upon which it is 
placed, and weighted with a couple of bars of lead on either side. 
To it are riveted two arms, say 2 feet 6 inches long and 2 feet apart, 
forked at the ends which, reaching into the lead, receive the wood ; 
they are connected half-way down by a cross-piece of flat iron. 
If the lead is to be poled with a billet of green wood, this and the 
crutch are best put in place before the kettle is filled, as thus the 
tedious work of depressing a green stick of wood into hot lead is 
avoided. It is much simpler and just as effective to introduce 
steam near the bottom of the kettle by means of a one-inch pipe 
bent to the form of the kettle and passing downward through the 
lead. 

The moulding of the lead from the merchant-kettle in American 
refining-works is almost always done with the Steitz siphon. 

In some of the older European works, where the height neces- 
sary for the siphon cannot be had without rebuilding the entire 
plant, the Rosing lead-pump* has come into universal use. 

Two moulding apparatus are shown in Figures 240-243. In 
Figures 240, 241, a represents the longer arm of the siphon, made 
of lj-inch pipe, with the cast-iron stop-cock b. At the lower end 
it is joined by two elbows, with a nipple intervening, to the swing- 
ing pipe b. This can be turned down around the centre c, and 
when in that position it can be moved in a horizontal circle, having 
its centre at d. The siphon is filled just like the one shown in Figure 
235, with the exception that the handling is done with the swing- 
ing pipe b, instead of with a pair of tongs. As in moulding, the 
swinging pipe has to describe nearly a semi-circle, the ordinary 
way of keeping the siphon in place by weighting with a couple of 
bars of lead is not sufficient. For this purpose two iron hoops 
about 2 feet 3 inches apart pass around the brick-work of the 

* Engineering and Mining Journal, November 28, 1885, and Berg- unci 
Huttenm. Ztg., 1889, p. 262. 
23 



PARKE 'S' PROCESS. 347 

kettle. The two ends of a hoop are bent, as shown in Figure 240, 
and tied by a bolt. The vertical arm of the siphon g is held in 
place between the two hoop-ends, the fixed bolt / and the movable 
one h. 

Another arrangement for moulding is shown in Figures 242- 
243. i represents the longer arm of the siphon with the stop-cock 
j; it discharges the lead into a 3-inch pipe k, closed at the bottom. 
This has two trunnions m, which swing in the bearings n. They 
are joined to a pivot rotating in the socket o. The pipe h is 
connected with the swinging arm p f which is moved with the 
handle q over the moulds. At some works the bottom of pipe k is 
closed with a cap having a socket, into which the pivot, fastened to 
the floor, fits loosely. Thus a number of slight variations in 
detail are found at different works. 

§ 110. Labor, Fuel, Output Of Lead.— In many refineries 
the working of the softening furnace, desilverizing-kettle, and 
refining furnace is given in contract as a whole, instead of having 
separate men working by the day at each furnace. Three men 
with a good head man can, if the work is well systematized, do 
everything that is necessary with a 30-ton plant. If the lead is 
moulded straight from the refining furnace, it is possible for them 
to attend to it also, but this is almost to overwork them. There- 
fore it has become the general custom to store the lead in the 
merchant-kettle and to mould from it. A separate contract is 
made for the moulding and loading into cars. This is often taken 
by a set of men who unload, weigh, and sample the base bullion, 
and deliver it at the softening furnaces. They also move the base 
bullion and lead produced at the blast-furnaces or the different 
reverberatory furnaces. By having a good head man for every 
desilverizing-plant and one or two contractors for handling the 
raw material, the by-products and the market-lead, the labor in the 
refinery becomes very much simplified and cheapened. 

The fuel consumed in softening, desilverizing, liquating, refin- 
ing, and moulding is about 330 pounds of soft coal per ton of 
base bullion. 

The amount of lead recovered in the form of market-lead 
varies somewhat according to the purity of the base bullion. It is 
about 80 per cent, of the bullion charged, or 88 per cent, of the 
softened lead in the kettle. 

§ in. Treatment of Zinc-Crusts.— The working of the zinc- 
crust has been and still is the weak point of Parkes' process. Many 



348 DESILVERIZATION OF BASE BULLION. 

methods 1 have been tried, but only few survive. They are all based 
on the volatility of the zinc and the readiness with which it is 
oxidized. Two only will be discussed. 

§ 112. Flach's Process. — This consists in smelting the zinc- 
crust in a blast furnace with a large percentage of slag, some 
matte and fluxes, the slag aimed at being ferruginous and low in 
silica, and the pressure of the blast not exceeding £ inch mercury. 
The zinc is partly taken up by the slag and matte ; a large part of 
it passes off into the dust-chambers ; the resulting rich bullion 
retains but little. If the crust contains any copper, it is taken up 
by the matte. In smelting zinc-crust with 120 per cent, of slag, 
30 per cent, of matte, 11 per cent, of fuel, and puddle-cinder as 
iron flux, in a small 36-inch circular blast furnace w r ith an Arent's 
siphon-tap, the writer found that after two days' running, a mushy 
substance collected on top of the lead, that refused to be taken up 
either by slag or matte. In order to keep up a good communi- 
cation with the lead in the crucible, the soft mush had to be 
repeatedly removed from the front of the furnace. It seems, 
therefore, that if in exceptional cases it should be necessary to 
smelt zinc-crust, it would be advisable to fill up part of the cruci- 
ble with brasque and to tap the lead and matte, when the* mush 
will be carried out of the furnace. 

While there is no doubt that smelting the zinc-crust in the blast- 
furnace furnishes quickly a rich bullion to be cupelled, the process 
has been abandoned as a regular method of treating the crusts, as 
none of the zinc is recovered, and the great losses of silver and lead 
are made up only in a very small degree by resmelting the zinc- 
bearing by-products. 

§113. Distillation of Zinc-Crusts. 2 — This process, first used 
by Parkes, has undergone many improvements and has become 
the one universally used in the United States since Balbach first 
used graphite retorts. The method, therefore, often bears his 
name. 

§ 114. Furnaces. — A furnace, to be suited for the process, 
must permit the raising and sustaining of a high temperature, and 
at the same time be of such a construction that a broken retort can 
be readily exchanged, and the rich bullion that has run out quickly 

JKerl, "Grundriss der Metallhuttenkunde," 1881, p. 314. Roswag 
" Desargentation de plomb," 1884, p. 296. 

2 Eilers, Trans. A.I. H.E., iii.. p. 314. Rising, Zeitschrift fur Berg-, Hut- 
ten-, und Salinen-Wesen., xxxiv., p. 91; Berg- und Huttenm. Ztg., 1886, p. 421. 



PARKE 'S' PROCESS. 



349 



and completely collected. Of the different forms three may be 
cited. 




JF-ig. ^44. 



1. The Faber die Faur Furnace (Figures 244-248).— This is 
a crucible furnace of cubical form, built into cast-iron frame- work 
that swings on trunnions, in order that the furnace may be turned 
over and the contents of the retort emptied. The furnace is closed 




Fig. 245. 



at the top by an arched roof, which usually has one opening, the 
charging opening for the coke ; the products of combustion pass 
off through a flue, which is generally placed at the back, as in the 



350 



DESILVERIZATION OF BASE BULLION. 



drawing, but sometimes at the side and occasionally in the roof. 
At the front is an opening for the neck of the retort. The bottom 
is formed by two sets of rectangular wrought-iron grate-bars placed 




Fig. 24 6. 



on edge. The retort rests on a small brick pillar, which is sup- 
ported either by a brick arch (as in the figure) or by an iron plate 
running from front to back, and protected from the heat by two 




Fig. 247. 

courses of fire-brick. The rotation of the furnace is effected by a 
worm-gear, sometimes simply by means of a lever. The furnace 
in its original form, as represented by the drawing, was about 4J- feet 
Cube, and was intended for a pear-shaped retort holding 250 pounds 
of liquated crust. It was lined with a full course of brick, except- 



PARKE 'S' PROCESS. 



351 



ing at the front, where the brickwork was 4J inches thick. At 
present the retorts, while retaining their original form, are made 
larger and thinner ; they hold as much as 1,000 pounds of crust ; 
the furnace has retained its original size, but is lined on all sides 
with a half course of brick. The old retorts were made of raw 
and burnt clay mixed with about 25 per cent, of graphite to pro- 
tect the clay from the corrosive action of the lead, and were very- 




Fig. 2-48. 

thick at the bottom ; at present they are made of graphite to 
which sufficient clay (perhaps 50 per cent.) has been added to give 
strength and stability to the retort. A 1,000- pound retort is 36 
inches high, 8 wide at the neck, 18 at the belly, and 13 at the 
bottom. It is 1J inches thick at the neck, and increases to 2 
inches at the bottom. 

The furnace has replaced most other furnaces since the patents 
of the inventor expired, proving the general favor it has won. A 
set of retorting-furnaces is arranged in two ways. They are 
placed on either side of a horizontal flue in such a way that the 
openings into it shall not be opposite each other, or they are built 
around a central stack, say eight in number, each flue extending 
into the stack and then continuing upward for a few feet. Both 
arrangements are so chosen as to avoid any obstruction of the 
draught. 

2. The Tatham Furnace (Figures 249-251). — This is a station- 
ary crucible furnace. The retort, which holds 500 pounds of zinc- 



352 



DESILVERIZATION OF BASE BULLION. 



crust, has the usual inclined position. It rests on a support at the 
back and protrudes at the front through a small arched opening. 
The top of the furnace is covered with a clay tile ; the products of 
combustion pass off through three small openings on one side, 
leading into a flue that is common to two furnaces, and terminates 
in the main flue leading to the stack. Each flue can be shut off by 

Fig. 249. 
SECTION ON THE LINE A B. 



TATHAM 
RETORTING-FURNACE. 




FRONT ELEVATION 
C 



Fig. 251. 

SECTION ON THE LINE CD. 




a damper. The bottom of the furnace, inclining from the back to 
the front, is made of brasque. It has a ridge in the centre and is 
elevated at the sides in order that any lead may run directly 
through the two gutters towards the front and out of the furnace. 
The coke is fed at the top, and the ashes and clinkers are removed 
at the bottom through the large opening in the front. Both front 
and back have stoking-holes. The admission of air is regulated 
by openings which are closed by bricks. The rich lead is tapped 
from the bottom of the retort and the residue raked out through 
the neck. The tap-hole is bored 1£ inches away from the side- 



PARKES' PROCESS. 353 

wall. The entire contents of the retort can, however, be removed 
through the neck, as is done with most stationary retorts. The 
ladle used for this purpose is a 6-inch piece of 3-inch gas-pipe 
closed at one end, and having an iron rod 5 feet long riveted to the 
other. It is good for six charges. Steitz ! several years ago con- 
structed a siphon to discharge the contents of a stationary retort. 
Sometimes it worked well and again it did not. As no reliance 
could be placed on it, all attempts at using it have been aban- 
doned. 

This furnace has replaced the Brodie furnace 2 at the Delaware 
Lead Works, which had two retorts, one above the other. The 
results obtained with the Tatham furnace by the writer have been 
very satisfactory. 

3. Other Furnaces. — In addition to the furnaces described, 
which are heated with coke, a few furnaces are in use that are 
constructed like a reverberatory furnace for the use of bituminous 
coal. The fireplace may be built on the side of the retort, at the 
back, or at the front, the aim always being to expose the retort as 
much as possible to the full action of the flame. The writer has 
worked a retort fired from the side with good results; retorts fired 
from the back appear to be satisfactory, while firing from the front 
has not proved so effective. The reason that retorts heated with 
bituminous coal have in so many cases given way to those using coke 
is because they require so much care to keep the temperature uni- 
formly at a white heat, which is absolutely necessary if the retort- 
ing is to be finished in the required time. Further, the facility 
with which the hot contents of a tilting-retort can be discharged has 
probably influenced the discarding of stationary retorts. The use 
of gaseous fuel has not been successful here as it has in Germany ; 3 
coal-oil is being experimented with at present, and promises very 
satisfactory results. 

§ 115. Condensers. — The condensers used for collecting the 
zinc differ ^ery much in form and material. Some are simply old 
retorts ; others are plumbago crucibles (diameter at bottom, 7 
inches ; 'at top, 11 inches ; height at front, 18 inches ; at back, 22- 
inches ; thickness, 1 inch). The former are supported by a spe- 
cially constructed buggy ; the latter rest on tripods, which also hold 
the receivers for the distilled zinc. Again, they are made of cast 
iron having the form of a truncated cone. One condenser of this. 

1 Egleston, "Silver, Gold, and Mercury," New York, 1887, vol. i., p. 102.. 

2 Trems. A. I. M. E., iii., p. 324. 3 Rosing, Loc. ciL 



354 DESILVERIZATION OF BASE BULLION. 

class is 2 feet long, and has handles on either side, by which it is 
suspended on two hooks from the iron frame of the furnace. A 
conical condenser is also made of clay, being about 3 feet long and 
supported by a tripod. Another form, finally, is that of a sheet- 
iron cylinder lined with specially moulded fire-bricks. At the base 
it has, in addition to the tapping-hole, two small pivots around 
which a thin chain passes and is hooked to the frame of the fur- 
nace, thus supporting the condenser. 

The condensed zinc is rarely allowed to run off continuously. 
Usually there is a tap-hole closed by a clay plug or a conical piece 
of coke, and the condensed zinc is discharged into a suitable mould 
only three or four hours after the distillation begins. It has been 
found that more zinc will collect in the condenser if it contains 
some liquid zinc. Most condensers have a second opening on the 
upper side for fumes to pass through while the distillation is going 
on. They go into a sheet-iron pipe leading to the main flue. It 
has always been considered essential to have this second opening 
if the distillation is to proceed in a satisfactory way, but at some 
works using an old retort as condenser the second opening has been 
dispensed with, the condenser being put in place only when the 
distillation begins, and not, as is customary, when the retort is well 
filled. 

§ 116. The Method of Working. — The method of working is 
about the same whichever furnace is used. When a new retort 
has been put in place, it is carefully warmed and brought up to a 
dull red heat. It is then ready to be charged. The zinc-crust, 
mixed with from 1 to 3 per cent, of charcoal, is brought in an iron 
wheel-barrow in front of the retort and charged with a trough- 
shaped scoop, filled on the wheel-barrow with a small shovel. The 
temperature is then quickly raised to a white heat, the crust softens 
and sinks in the retort after half an hour, is pushed down, more 
crust is added, and this is repeated till the retort is completely filled. 
The larger rim of the condenser receives a heavy lute of clay, is 
passed over the neck of the retort, and is made to adhere to the front 
wall of the furnace ; the lower end rests on its support. About an 
hour after charging, blue powder and then metallic zinc begin to 
collect in the condenser. The distillation is finished after from six 
to eight hours, according to the size of the charge, the percentage 
of zinc in the crust, and the draught of the furnace. Too much 
emphasis cannot be laid upon the draught. The slightest obstruc- 
tion means a failure in distilling off the zinc in the required time, 



PARKE 'S' PROCESS. 355 

and many consequent disturbances, inconveniences, and losses. 
The main points to be looked after during the operation are the 
quick raising of the temperature, and the keeping it high. The 
higher the temperature, the more rapid will be the distillation, the 
better the output of metallic zinc, and the lower the percentage of 
zinc remaining in the rich lead. If the temperature is lowered, 
blue powder forms, and some of the crust floating on the lead may 
harden and be suddenly broken by the zinc-vapors developed be- 
neath the crust, when the heat is raised again, and thus cause an 
explosion. The effect of this is to loosen the condenser from the 
neck of the retort. At most works it is the practice to introduce 
at certain intervals, through the upper opening in the condenser, an 
iron rod, free the neck of the retort with it from oxidized zinc, 
and then stir up the crust floating on the lead. The saying that 
the condenser ought to smoke well if the distillation is to go on 
satisfactorily, is a mistaken one when a condenser with only one 
opening — the tap-hole, which is kept closed — is giving excellent 
results. 

When the zinc ceases to collect in the condenser the distillation 
is finished. The last zinc is then removed and the condenser taken 
off and scraped clean. In the meantime the vapors in the retort 
pass off into the air. It is assisted by throwing a few chips 
of wood into the retort. It has now to be emptied. With a 
tilting-furnace, a slag-pot, lined with brick to prevent the hot 
metal from cracking the cold bottom, is wheeled in front of the fur- 
nace and the lead emptied into it. While the lead is running out 
the sample for assay is taken. It is then allowed to cool and, after 
the removal of the dross from the surface, is ladled into small 
moulds, so as to obtain bars of suitable size for the subsequent 
treatment in the English cupelling-furnace. The residue in the 
retort, consisting of slag and charcoal, is removed with an iron 
scraper. It is essential for the life of the retort that it be well 
cleaned after each distillation. The retort is now raised, some 
fine charcoal is thrown into it to prevent the oxidation of globules 
of lead adhering to the sides, it is then turned back to its normal 
position, and is ready to be charged again. The grate-bars 
are cleaned, clinkers adhering to the furnace-walls or sides of the 
retort are removed, fresh coke is added, and all is then ready for 
the next operation. The time required for discharging and refilling 
is about twenty minutes. In scraping, care must be taken to do it 
gently, so as not to wear off the lower side of the retort. To avoid 



356 DESILVERIZATION OF BASE BULLION. 

this, as well as to change the line of contact with the surface of 
the lead, the position of the retort is sometimes changed, after it 
has worked about twenty charges, by turning it 180 degrees. 

With the Tatham furnace the operations are similar, except that 
the lead is tapped from the bottom, and only the residue raked out 
through the neck. 

§ 117. Tools. — The tools required by one man in retorting are : 
2 scrapers (6 feet long, -|-inch round iron flattened out at one end to 
the width of 3 inches and bent up 3 inches) to stir the contents of 
the retort after distilling and to rake out the residue after tilting); 
2 pokers (5 feet long, of |-inch iron) to remove the clinkers from 
the grate ; 1 bar (8 feet long, j-inch steel) to break off clinkers 
from the walls ; 1 scoop to feed the coke ; 1 ladle (6 inches 
in diameter and 3 inches deep, with a 4-foot handle) ; 2 slag-pots 
lined with brick ; a wheel-barrow to receive the zinc-crust ; 1 
scoop and shovel to charge the retort ; 2 moulds for every retort 
to receive the zinc ; 10 bullion-moulds. 

§ 118. Results. — The weight of the charge is from 800 to 1000 
pounds of zinc-crust, and it takes from six to eight hours for one 
operation ; less time is required in winter than in summer. The 
crust yields from 70 to 80 per cent, of lead containing from 4 to 
10 per cent, of silver and from 0.75 to 1.50 per cent, of zinc. 
The zinc recovered in the form of metal is about 10 per cent, of 
the weight of the crust, and that in the form of blue powder about 
1 per cent. Of all the zinc required for desilverizing, over 60 per 
cent, is recovered to be used again in the kettles. The quantity of 
dross varies a great deal ; from 5 to 8 per cent, of the weight of 
the crust is a fair figure. A retort lasts now forty charges ; 
formerly twenty-five charges was considered a good average. For 
every ton of crust 1,100 pounds of coke are required. To use a 
good quality of coke of uniform size (egg-coke), although more 
expensive at first, is cheaper in the end, when the life of the retort 
and the better results obtained are considered. One man attends 
to from two to four furnaces in a twelve-hour shift. 

§ 119. Roesing's Suggested Improvements.— Roesing ' pro- 
poses to facilitate the distilling of zinc-crust in a very novel way. 
He heats the retort, having a basic lining in a tilting-furnace. 
When the crust has become soft, he introduces some fuel to drive 
out all the air, and pours pig-iron, heated above its melting point, 
into the retort, which volatilizes the zinc to be caught in a con- 
1 Berg-und BUttenm. Ztg., 1890, p. 369. 



PARKE 'S' PROCESS. 357 

denser. By tilting the retort, first the iron is poured off to be 
used over again and then the rich lead. 

§ 120. Comparison of the Two Processes. — In comparing the 

smelting of the zinc-crust in the blast furnace with the distillation, 
it is clear that if the cost of distilling is made up by the amount of 
zinc recovered, it is preferable to smelting, on account of the greater 
loss of silver incurred in the latter method. General practice has 
shown that there is even a margin of profit in distilling the silver- 
crust, and hence the smelting has been almost entirely abolished. 
There is one case when it may be doubtful, which is more profitable. 
The gold-crust is very heavy, compared with its low contents in 
silver. In retorting, very little metallic zinc is recovered, most of 
it being obtained in the form of blue powder ; the retort-bullion 
assays from 400 to 800 ounces silver per ton and is rich in copper. 
Some works, therefore, prefer to smelt this crust with the addition 
of matte, obtaining thus a bullion which is free from copper and 
is easily cupelled. 

§ 121. Treatment of By-Products. — This is a very important 
part of refining. It requires to be carried on simultaneously with 
the main operations, so that the by-products may be disposed of at 
once and not allowed to accumulate. 

§ 122. Softening Furnace Dross and Skimmings. — These are 
usually first liquated in a reverberatory furnace, in order to extract 
as much lead and silver as possible. One furnace of reasonable 
dimensions, if kept constantly in use, is able to liquate both prod- 
ucts of a plant treating 100 tons base bullion a day. While a 
furnace for liquating the dross ought properly to have a some- 
what different construction from the one used to treat the anti- 
mony skimmings, it is found more economical to have one furnace 
for the two operations. The reverberatory (for example, one with 
a hearth 8 by 12 feet and 10 inches deep, and having two working- 
doors on one side only) is built just like a softening furnace. It 
also has a water jacket to counteract the corroding influence of 
the antimony skimmings when they are being melted down. The 
two tap-holes of such a furnace are placed in one of the sides near 
the centre. 

In liquating dross, the furnace is charged and heated until the 
lead that has liquated forms a bath on which the dry dross 
floats, the operation being continued until sufficient dry dross has 
accumulated on the lead, that it may be raked out through the 
furnace-doer. The dross directly in contact with the lead is not 



358 DESILVERIZATION OF BASE BULLION. 

touched till towards the end. When the level of the lead comes 
near the furnace-door, enough is tapped into an outside kettle to 
leave a bath of lead for the charge next following to float on ; 
charging, heating, and drossing are continued until the batch of 
dross is worked up. The temperature is always kept low, in order 
that no dross may be taken up again by the lead. The liquated 
dross is smelted in the blast furnace with a sulphur-bearing ore or 
by-products (galena, matte low in copper) to form a matte, or to 
increase the percentage of the matte that has added to the charge. 
The lead goes to the softening furnace. 

In liquating the softening furnace skimmings, two objects are 
aimed at, the elimination of metallic lead and the desilverization 
of the skimmings. Both are accomplished at once by melting 
down the skimmings with a reducing [i.e., a smoky) flame in 
the reverberatory furnace. All the lead carried off from the 
softening furnace will collect on the bottom of the hearth, and, 
from the melted skimmings floating on top, part of the lead 
oxide, converted into metal by the reducing flame, will, while 
descending through the skimmings, desilverize these and carry the 
silver with it into the metal-bath below. If necessary, some fine 
coal is spread over the charge. When the furnace is filled and 
everything well melted, the liquated skimmings are tapped from 
the upper tap-hole into slag-pots, and the lead from the lower tap- 
hole into a kettle, whence it is ladled into bars to go back to the 
softening furnace. 

Another and better way is to have only one tap-hole, and to tap 
the entire contents of the furnace into a kettle that has an over- 
flow spout. While the lead is flowing into the kettle some can 
be ladled out, if necessary. As soon as the liquated skimmings 
appear, they will rise in the kettle and overflow through the spout 
into slag-pots placed beneath. When the flow ceases, the skim- 
mings still floating on the lead will soon harden and are then 
removed ; the lead in the kettle is moulded by hand into bars and 
goes to the softening furnace. 

When the cake of liquated skimmings is rolled out from the 
slag-pot and broken, there will be found two products : liquated 
skimmings assaying three ounces silver or less to the ton, form- 
ing the bulk of the cone, and at the bottom a small cake of a 
silvery-white antimonial speise of lead and copper, which contains 
as much as 40 per cent, copper, and with 250-ounce bullion assays 
often as high as 300 ounces silver to the ton. Formerly it was- 



PARKES' PROCESS. 359* 

considered preferable to tap the skimmings or let them overflow 
on an iron plate, as they chilled quickly and were easily broken to 
pieces, but they ran higher in silver. 

The liquated skimmings are smelted in a small (36-inch) blast 
furnace, with about 100 per cent, of slag and 11 per cent of fuel, 
care being taken to avoid a slag that is very ferruginous, as speise 
might otherwise form. In order to reduce the loss of antimony 
by volatilization (which is considerable), non-argentiferous galena 
is added to the charge. The sulphur acts as a reducing agent, 
doing thus to some extent the work of carbon or carbonic oxide. 
The amount of galena added is so regulated that no matte shall 
form ; it is from 13 to 28 per cent, of the weight of the skim- 
mings. While this addition reduces the percentage of antimony 
in the resulting hard lead, this is of no consequence, as, if hard 
lead assays from 14 to 20 per cent, antimony (a common figure), this 
is paid for at the same rate as when the lead contains from 7 to 10- 
per cent. 

The resulting slag is always liable to be rich in lead and 
to contain considerable antimony. Smelting the slag all again 
in the ore blast-furnace would bring the antimony back to the 
base bullion, and with every smelting of hard lead, fresh ore- 
slag would be required. Therefore, a certain amount of slag is 
kept apart and run over and over again, and only the small 
excess, produced in every run from coke-ash and fluxes given 
back to the ore-charge ; the rest is saved for the next antimony 
run. 

Another way to treat the softening skimmings is to smelt them, 
without liquating, in the blast furnace and to make a hard bullion. 
This is charged into the softening furnace with soft lead poor in 
silver (say from the Mississippi Valley), in the proportion of one 
hard lead to four soft lead. The amount of silver in the original 
hard bullion is reduced, the softening facilitated, and the resulting 
skimmings are low enough in silver to be smelted directly into 
marketable hard lead, which, however, ordinarily runs as high as 
6 ounces silver per ton. 

The grade and grain of the hard lead are much improved by 
poling it in a kettle for a few hours or liquating it in a reverber- 
atory furnace at a low temperature. The reason for this is that 
the copper remaining in the softening skimmings enters the hard 
lead on being smelted. The dross from liquating the crude hard 
lead has been found to contain as much as 40 per cent, copper, 



360 DESILVERIZATION OF BASE BULLION. 

but generally contains about 10 per cent. The hard lead is 
usually moulded by ladling from the kettle where it has been poled, 
or from the one into which it ran from the liquating furnace. 
The surface of the refined hard lead, when liquid, soon becomes 
covered with dross, and in order to obtain clean bars without 
being forced to waste much lead by skimming, it is advisable 
to place a wrought-iron ring on the lead just large enough 
for the ladle to pass through and to dip the lead from the 
ring. Thus only a very small surface will have to be kept 
bright. 

Another way of treating the dross and skimmings has been in- 
troduced at some works within a few years, by which the copper 
is eliminated from the skimmings before they are smelted in the 
blast furnace. It is to smelt dross and skimmings together in a 
reverberatory furnace with some galena free from silver. The 
result is base bullion, a copper-lead matte free from iron, and 
refined skimmings. The skimmings are discharged through the 
upper, matte and lead through the lower tap-hole. The skimmings 
are low enough in silver to be smelted for hard lead, and being free 
from copper and other impurities can be ladled directly from the 
lead-well into marketable bars. Thus the common operation of 
liquating or poling hard lead from the blast-furnace is made unne- 
cessary. The bullion goes back to the softening furnace. The 
matte is converted in a cupelling furnace with some siliceous 
material into copper bottoms, 60 per cent, copper-matte and slag, 
the bottoms collecting the gold. The slag runs off into an iron pot 
and, when the test is full of metal and matte, is tapped into 
moulds, placed on a truck beneath the test, by drilling with an 
augur a hole into the breast. The tap-hole is again closed from 
the inside by inserting a clay stopper at the back through the 
tuyere-hole ; the furnace is then ready for another charge. When 
the moulds are cold, the bottoms are separated from the matte, 
and this is converted in another cupelling furnace into metallic 
copper to be used in precipitating silver from silver sulphate solu- 
tion obtained in parting dore silver by means of sulphuric acid. 
The advantage for refining works in concentrating the copper 
obtained from base bullion in this way lies in the fact that copper 
refiners pay for only 93 per cent, of the silver contained in the 
matte, making no allowance for the lead, and charge two and a 
half times as much for desilverizing matte as they do for metallic 
copper, or $40 per ton as against $100. 



PARKES' PROCESS. 361 

Analyses of Hard Lead and Hard Lead Dross. 





Lautenthal. 


Clausthal. 


Pribram. 


Before Poling. 


After Poling. 


Dross from liquat- 
ing Hard Lead. 


Sb 


12.753 

'i.'sei 

85.291 
0.0035 

trace 
trace 


15.390 

o.ihk 

84.650 
0.0030 

trace 
trace 


38.763 

37! 643 
22.962 

(S0\240) 
0.139 
0.232 


24.270 
0.109 
0.160 

74.886 
0.009 
0.015 
0.018 
0.009 
0.524 


As 


Cu 


Pb 


As 


Ni 


Fe 


Zn 


Sn 




Reference 


i 


3 


3 


4 



§ 123. Tin Skimmings. — Tin ore occurs so rarely with lead- 
silver ores that it is only an exceptional case when tin skimmings 
are formed in softening base bullion. At Freiberg (Saxony) the 
tin of the ores becomes concentrated to some extent in the base 
bullion produced. The skimmings that rise to the surface in the 
softening furnace after the dross has been removed have the fol- 
lowing composition : 5 

PbO, 70.35 
Sn0 2 , 12.53 
Sb 2 5 , 12.50 
As 2 5 , 4.73 
CuO, 0.61 

and contain 72.9 ounces silver to the ton. Plattner has introduced 
a method of desilverizing these skimmings and concentrating the 
tin in a marketable alloy which contains 

Sn, 33 
Sb, 14 

As, 1 

the rest being lead. Details with analyses of intermediary prod- 
ucts are given in the reference. 

1 Private notes. 2 Ibid. 

9 Berg- und Huttenm. Ztg., 1870, p. 169. 

* Oesterreichisches Jahrbuch, xxxix., p. 64. 

6 Plattner, Berg- und Huttenm. Ztg., 1883, p. 417. 



362 DES1LVERIZATI0N OF BASE BULLION. 

§ 124. Kettle DroSS. — The impurities contained in this dross 
are very few. It consists principally of lead oxides mixed with 
metallic lead, and is usually put back into the softening furnace 
with the next charge after the furnace dross has been drawn off. 
This is the best way of disposing of it. 

§ 125. Refining Skimmings and Polings.— These are best* 
treated in a reverberatory furnace that is used for no other pur- 
pose, except perhaps for liquating hard lead, in which case the 
lead resulting from the refining skimmings is sold as second-class 
lead on account of the copper it has taken up. If the reverbera- 
tory is used only for reducing refining skimmings, the resulting 
lead can be worked in with the regular refining charges and corrod- 
ing-lead produced. The hearth of a furnace suited for this pur- 
pose may be 10 by 5 feet and 9 inches deep, built of fire-brick 
enclosed in a wrought-iron pan. It will have an inclination of 3 
inches from the bridge to the flue, where the main working-door is 
placed ; a second door is in the middle of one of the sides to intro- 
duce and distribute the charge. The tap-hole below the flue dis- 
charges the lead into a small spherical kettle having a fire-place 
beneath. The English (§ 43) or Silesian (§ 44) reverberatory also 
serves the purpose very well. 

The mode of operation is as follows. The hearth of the fur- 
nace, dark-red from a previous charge, is covered with a 2-inch 
layer of fine coal to protect it from the corroding effect of the 
skimmings. It is then filled with skimmings mixed with 10 per 
cent, of fine coal, leaving room for the gases to pass. Sometimes 
the charge reaches only to the working-doors, being renewed 
from time to time as it shrinks. As the charge heats, lead will 
flow into the sump, whence it is tapped at intervals. The fire is 
gradually urged ; when no more lead flows the charge is rabbled, 
and, when exhausted, drawn. 

Four tons of skimmings, yielding about 60 per cent., are worked 
in twelve hours, two men attending the furnace. 

The residue, which contains lead, zinc, antimony, perhaps some 
arsenic, and coal-ashes, is added to the smelting charge of liquated 
softening skimmings. 

The refining skimmings and polings are sometimes, though 
rarely, smelted in the ore blast-furnace with the original softening 
skimmings to reduce the silver-contents of the latter. They are 
also sometimes charged into the softening furnace after the dross 
has been removed, in order to oxidize the arsenic and antimony of 



PARKES' PROCESS. 363 

the bullion, but some zinc enters the hard lead when the skimmings 
come to be smelted in the blast-furnace, making it unfit for some 
of its uses, as for example in acid works. 

Owing to the scarcity of lead ores now prevailing, nearly all 
refiners use up their refining skimmings for lead-flux in smelting 
dry argentiferous ores. 

§ 126. Rich Lead and Metallic Zinc. — The former goes to 
the cupelling furnace (§ 133) ; the latter is used in desilverizing. 
It always contains a few ounces of silver. 

§ 127. Retort Dross and Bine Powder. — The retort dross is 
very rich in silver, which must be extracted quickly. The dross, 
if formed in small quantities, is worked off in the regular cupelling 
a little at a time, or at some works it is charged back into the 
retort. If there is too much for this, it is allowed to accumulate, 
and is scorified on the bath of lead free from or low in silver, with 
which a new test is usually charged, in the cupelling furnace. 
Sometimes the retort dross is added to the softening furnace, 
after its own dross has been removed, that the silver may be 
taken up by the lead, and the impurities oxidized and taken up 
by the skimmings. It is not often that the dross is added to the 
charge of the bullion blast-furnace. 

The blue powder, consisting of a mixture of finely divided 
metallic zinc and zinc oxide, always contains some silver, say 
from 4 to 5 ounces per ton. It is not readily disposed of. If dis- 
tilled by itself with the addition of charcoal, it yields from 33 to 
55 per cent, of zinc. At some works it is fed back to the retort 
with the following charge of zinc-crust ; at others it is added with 
the first zinc to the desilverizing-kettle, that the metallic zinc may be 
taken up. If this is done, 50 pounds of blue powder are charged at 
a time into the kettle before the lead is tapped into it from the 
softening furnace. It is stirred into the bullion while this is rising 
in the kettle. It does not remove much silver from the kettle, but 
serves rather to saturate the lead with zinc and to assist in remov- 
ing the gold and copper. Sometimes the blue powder is screened 
to remove all coarse particles and is then sold to zinc-works. 

§ 128. Litharge is reduced in a reverberatory furnace or goes 
to the bullion blast furnace ; sometimes it is added to the soften- 
ing furnace after drossing. While the reduction of litharge in the 
reverberatory furnace is preferable from a purely metallurgical 
point of view, the blast-furnace is in more common use, as refining 
works are thus often enabled to smelt dry silver ores at their own 



364 DESILVERIZATION OF BASE BULLION. 

price, as the Western smelters do not compete for them on account 
of the usual scarcity of lead ores. It is important for a refinery to 
smelt some ore in treating by-products in order to make new slag, 
and not be forced to smelt old slag over and over, which is expen- 
sive and causes losses in lead and silver, as these slags become grad- 
ually rich in zinc. 

§ 129. Old Retorts, Cupel-Bottoms, etc.— These are added to 
the charges of the bullion blast-furnace. 

§ 130. Table of Desilverization. —The following table gives 
a summary of the work done in an American desilverizing works, 
as shown in detail in the preceding pages : 

1. Base Bullion : softened in reverberatory furnace yields soft 
bullion 2 , softening dross 7 , softening skimmings 8 . 

2. Soft Bullion : desilverized in kettle yields kettle-dross (back 
to 1), gold-crust 3 , silver-crust 11 , desilverized lead 12 . 

3. Gold-Crust : liquated yields liquated gold-crust 4 , and gold- 
bearing base bullion a . 

4. Liquated Gold-Crust : retorted yields retort dross 15 , blue 
powder 14 , zinc 3 , retort (dore) bullion 5 . 

5. Retort (Dore) Bullion : cupelled yields dore silver 6 , 
litharge 16 , cupel-bottom 16 . 

6. Dore Silver : sold, or parted by sulphuric acid or electrolysis, 
yields gold, silver, blue or green vitriol, or precipitated copper. 

7. Softening Dross : liquated in reverberatory I. yields base 
bullion 3 , an d liquated dross 16 . 

8. Softening Skimmings : liquated in reverberatory I. yields 
base bullion *, and liquated skimmings 9 . 

9. Liquated Skimmings : smelted in 17 yield hard lead 10 and 
slag". 

10. Hard Lead : poled in kettle or liquated in reverberatory II. 
yields market hard lead and dross 16 . 

11. Silver-Crust : goes to 3, 4, and 5 in separate apparatus, 
yielding fine silver and by-products treated like 4 and 5. 

12. Desilverized Lead : refined in reverberatory yields corrod- 
ing-lead and refining skimmings 13 . 

13. Refining Skimmings : reduced in reverberatory II. yield 
lead (either to 12 or sold) and dross 17 , or smelted with ore in 16. 

14. Retort Dross : either fed on cupel 6 , goes back to retort 4 , 
softening furnace \ or bullion blast-furnace 16 . 

15. Blue Powder : goes either to 2 (as first zinc), or back to 
retort *, or screened and sold. 



PARKE 'S' PROCESS. 365 

16. Bullion Blast-Furnace. 

17. Hard-Lead Blast-Furnace. 

§ 131. General Remarks. — A few general remarks regarding 
Parkes' process are in place. 

The output of silver with average base bullion as it comes 
from Western smelting works (running from 150 to 250 ounces 
silver, and 0.5 ounces and less gold per ton, being pretty free from 
dross, but containing some arsenic and antimony) is never less than 
99.5 per cent., and generally there is a slight surplus of silver ; 
with gold it is from 98 to 100 per cent, (a surplus is rare) ; with 
lead from 99 to 99.5 per cent. 

A refinery makes its deduction for loss in treatment and for 
cost of working, the base bullion being delivered f. o. b. refinery. 
Some refining works sum up both items and make a general charge 
by the ounce of silver ; e. (/., one cent for every ounce of silver, 
the lead being paid on New York quotations and the freight to New 
York being deducted. Others make a working charge by the ton 
and a deduction for loss in silver ; e.g., the working charge varies 
from $10 to $14 per ton, and the reduction amounts to 2 or 3 
ounces silver per ton, the lead being paid as above. A number 
of other ways of settling between smelter and refiner are .in use, 
but the two quoted are the most common ones. 

No detailed statement can be made about the cost of refining* 
that would be generally applicable, as the single items vary greatly 
in different refineries. In a general way it may be said that the 
actual operating expenses, including interest and taxes, vary from 
$5 to $6 per ton of base bullion. If the general expenses, such 
as salaries, marketing, etc., be added, this figure rises to $8 and 
$9. If, finally, the loss in metal and incidentals is added to the 
last figure, giving the total final cost, this will be found to vary 
from $10 to $12 per ton. 

In a refinery where so many by-products result, it is essential to 
know how much silver, gold, and lead is contained in each of 
them. These quantities vary a great deal with the different kinds 
of bullion treated. It is therefore essential to be able to follow 
up the metals from the time they enter the refinery until they 
leave it. 

Below is reproduced the scheme of the Kettle-Book, the most 
important one of all the books. It shows the various products 
and the amount of fuel consumed : 



366 



DE8ILVERIZATI0N OF BASE BULLION. 



Kettle-Book.* 



189- 
Date. 


Lot. 


Name of 
Bullion. 


Charge. 


Gross 
Weight 


Assay : 
Ozs. per ton. 


Contents : 
Total Ounces. 


Month. 


Day. 


No. 




No. 


Lbs. 


Ag. 


Au. 


Ag. 


Au. 























Softening 
Dross. 


Softening 
Skimmings. 


Kettle Dross. 


Net Weight 
in Kettle. 


Assay : 
Ozs. per ton. 


Contents: 
Total Ounces. 


Lbs. 


Lbs. 


Lbs. 


Lbs. 


Ag- 


Au. 


Ag. 


Au. 



















Zinc Added: 
Lbs. for. 


Gold-Crust. 


Assay : 
Ozs. per ton. 


Contents : 
Total Ounces. 


Ag. 


Au. 


Lbs. 


Ag. 


Ag. 













Liquated 
Gold-Crust. 


Liquated 
Silver-Crust. 


Refining 
Skimmings. 


Refined 
Lead. 


Coal. 


Remarks. 


Lbs. 


Lbs. 


Lbs. 


Lbs. 


Tons. 

















It is to be noted that the " net weight in kettle," before add- 
ing the first zinc, is found by deducting the sura of weights of 
softening dross, skimmings, and kettle dross from the base bullion 
charged into the softening furnace. The " net weight in kettle " 

* Owing to the form of this volume, it has been necessary to put the head- 
ings in four rows, one beneath the other, while in the actual kettle-book, they 
simply run across the two opposite pages in one line. 



PARKES' PROCESS. 



367 



after adding the first zinc, is found in a similar way, by adding the 
pounds of zinc used for extracting the gold to the net weight 
before the first zinc addition was made and deducting the weight 
of the gold-crust from this sum. 

The principal other books kept in a refinery are : the retort- 
book, the cupellation-book, and the books for reverberatory furnace 
I. (liquating softening dross and skimmings) and II. (reducing 
refining skimmings and liquating hard lead), and special assay- 
books for the softening furnace, desilverizing-kettle, retorts, and 
cupelling furnaces. 

§ 132. Relative Advantages of Parkes' and Pattinson's Pro- 
cesses. — The many advantages 1 (§ 90) that Parkes' process has over 
Pattinson's have made it the desilverizing process used in the United 
States, there being only one refinery (at Eureka, Nevada), that 
desilverizes with the Luce-Rozan process. There is, however, one 
instance in which Pattinson is to be preferred to Parkes. It is 
when the base bullion is rich in bismuth. During the crystallization 
bismuth follows the liquid lead (§91). It is thus concentrated in the 
rich lead, and can be recovered when this is being cupelled. The 
bismuth-contents of the market-lead thus become very much lowered, 
although not quite removed. In Parkes' process the bismuth enters 
only to a small extent (§ 94) into the zinc-crust, with the result that 
the refined lead may become richer in bismuth than the original bul- 
lion. This difference is shown by the two following analyses by 
Hampe 2 from Lautenthal market-lead where Parkes' process replaced 
that of Pattinson. 



Lautenthal. 



Pb. 



Pattinson. 

Parkes.... 



Cu. 



Sb. 



As. 



Bi. 



Ag. 



Fe. 



Zn. 



.9662 0.015 0.010 jnone 0.0006 0.00*22 0.004 0.001 
.983139 0.001413 0.005698jnone 0.005487 0.000460 0.002289 0.000834 



Ni 



0.001 
0.000680 



An interesting combination of the two processes is found at 
Freiberg, 3 where the original base bullion, containing from 116 to 
233 ounces silver per ton, and from 0.02 to 0.06, rarely 0.16 per 
cent, bismuth, is concentrated by crystallization to a liquid lead 
with 0.17 per cent, bismuth, assaying 583 ounces silver per ton, 

1 Phillips, Engineering and Muring Journal, May 27, 1887. 

' Zeitschrift fur Berg-, Eutten-, und Salinen-Wesen in Preussen, xviii., p. 
195. 

• Engineering and Mining Journal, December 4, 1886 ; Berg- und Huttenm. 
Ztg., 1887, pp. 45, 192. 



368 DESILVERIZATION OF BASE BULLION. 

which is cupelled, while the crystals averaging 30 ounces silver 
per ton are desilverized with zinc, the bismuth in the market-lead 
not exceeding 0.02 per cent. 

Large shipments of base bullion with considerable quantities of 
bismuth are of rare occurrence in refining works. Occasional ones 
are worked in with other bullion that is free from, or at least low 
in 3 bismuth. 



CHAPTEE XI. 
CUPELLATION. 

§ 133. Introductory Remarks. — The process of cupellation 
has for its object the final separation of lead and silver. It con- 
sists in melting and heating in a reverberatory furnace argen- 
tiferous lead with access of air to the temperature at which 
litharge forms on its surface. This is run off and is in part ab- 
sorbed by the hearth, while the silver, having scarcely any affinity 
for oxygen, remains behind in the metallic state. The oxidation 
of the lead is principally effected through the action of the blast 
playing over its surface, but is also assisted by the litharge when 
formed, as this absorbs oxygen and gives it off again to the leadi 
and its impurities. The most important of these are copper, arsenic,, 
antimony, bismuth, silver, and gold. 

The bulk of the copper is removed with the dross, all the arsenic 
and antimony with the skimmings, just as in softening base bul- 
lion (§ 97). The copper remaining with the lead after drossing, is 
taken up only very gradually by the litharge. As it has less affin- 
ity for oxygen than lead, its oxidation must be caused not so much 
by the action of the air as by a large quantity of litharge acting 
on a small amount of copper. With reversed conditions cuprous 
oxide 1 oxidizes lead. The oxidizing action of cuprous oxide con- 
tained in the litharge seems to be the cause of the fact that cuprif- 
erous lead is cupelled quicker and with less loss in lead and silver 
than if the lead were free from copper. Kerl 2 states that in cupel- 
ling thirty tons argentiferous lead containing from J to 1 per cent, 
copper, the operation lasted twenty-four hours less than when 
copper was absent, and that the losses in metal were as 2:7. 

Bismuth is concentrated with the silver in the lead until towards 
the end of the process, and then greatly retards the progress of the 
work. It is finally oxidized and enters the litharge, giving this a 
greenish color, and is absorbed by the hearth-material, while th& 

1 Berthier, "Traite des essais," Liege, 1847, vol. ii., p. 572. 

2 "Grundriss der Metallhuttenkunde," 1881, p. 270. 

24 



3T0 DESILVERIZATION OF BASE BULLION. 

silver also retains some bismuth with great tenacity. If the bis- 
muth is to be recovered, the silver is concentrated only to a certain 
degree (say to 50 or 00 per cent.) in the lead, which is then cupelled 
in a separate furnace ; the saturated part of the hearth of this 
furnace and the litharge form the raw material for the extraction 
of bismuth in the wet way. 

Silver is always present in the litharge, probably in both forms, 
as oxide and as metal. Rose ' states that silver oxide begins to 
lose its oxygen at 250° C. ; according to Sainte Claire-Deville and 
Debray, 2 as well as Troost and Haute-Feuille 3 , the oxide appears 
to be able to exist at a very elevated temperature. Finally Wait 4 
dissolved from litharge containing 2.94 per cent, silver, by means 
of acetic acid, 18.67 and 19.25 per cent, of the silver present. As 
metallic silver is insoluble in acetic acid, the silver dissolved must 
have been present as oxide. If silver-bearing litharge remains in 
contact any length of time with metallic lead having little silver, 
it loses its silver. Thus at the beginning of the cupellation little 
silver is contained in the litharge. During the progress of the 
operation the lead becomes richer, more silver is liable to be 
oxidized, and less of it again reduced by the enriched metallic 
lead. Fine particles of argentiferous lead cannot be prevented 
from being carried away by the litharge. From both causes, 
therefore, the tenor in silver of the litharge will increase with the 
progress of the cupellation. 

Gold, finally, follows the silver in the cupellation, but none of 
it, or perhaps a trace, are found in the litharge. 

According to the general construction of the furnace, and the 
consequent mode of operating, cupellation is generally discussed 
under two heads : 

A. German Cupellation. 6 

B. English Cupellation. 

1 "Handbuch der analytischen Chemie," Leipsic, 1867, i., vol. 339. 

2 Graham-Otto-Michaelis, ' ; Anorganische Chemie," Brunswick, 1884, iii., p. 
985. 

' Ibid. 

4 Trans. A. I. M. E., xv., p. 463. 

6 The writer is fully aware that the German cupelling furnace is not much 
found in this country ; in fact, he knows of only one refinery that uses it. If, 
nevertheless, it is treated here in more detail than may seem necessary, the 
reason is that operations which it has in common with the English furnace can 
be more easily dealt with here than later. 



CUPELLATION. 371 

A. German Cupellation. 

§ 134. Characteristics. — The characteristics of this method are : 
a large reverberatory furnace with a fixed bed and a movable 
roof; the fact that the bullion to be cupelled is all charged at once, 
and the silver not refined in the same furnace where the cupellation 
was carried on. 

§ 135. The Furnace. — The furnace selected for illustration is 
the one in operation at Pribram 1 (Figures 252-255). It differs 
somewhat from the generally accepted circular form, 2 and is an 
improvement on it. Figure 252 shows the fire-place a at the right, 
and the flue b at the left, of the hearth. Figure 253 represents 
a horizontal section of one furnace and the fire-place a' of a 
second furnace, the furnaces being built in pairs. The products of 
combustion pass downward through four separate flues b, which 
unite in one main flue c, leading to the dust-chambers. In Figures 
254 and 255 are seen two vertical sections on the same line EF, 
Figure 254 representing the furnace before, Figure 255 after, 
tamping in the hearth. The furnace is built of common red brick, 
with the exception of the parts that are exposed to the flame, as 
indicated by the cross-hatching. In the upper part of the founda- 
tion and in the side-walls small channels d and e reaching out- 
ward are left open. These serve as drains for the moisture. At 
the back of the furnace are three openings J] through which the 
tuyere-pipes are introduced. At the front is the litharge-channel 
g, which can be closed by a sliding-door ; a cast-iron breast-plate 
h serves as support for the upper hearth i. The movable arched 
roof h rests on an L ring £, and is removed with a differential 
pulley suspended from a traveller. On the foundation is built a 
brick bottom m, the brick being set dry. Beneath its lowest 
point, just below the cavity n, is a cast-iron plate o, to prevent any 
leakage of metal through the drain d, should the working-hearth 
crack or be injured in any other way. The flues leading from the 
fire-place a to the hearth are shown in j. The fire-brick part of 
the furnace is encased in cast-iron plates that have openings cor- 
responding to the upper drains. The whole furnace is well bound 
together by buckstays and tie-rods. The fuel used is a mixture of 
bituminous coal and lignite ; the ash-pit is closed as the blast is 
introduced beneath the grate-bars. 

1 Oesterreichisches Jahrbuch, xxxix., p. 46. Private notes, 1890. 

2 Zeitschrift fur Berg-, Hutten-, und Salinen-Wesen in Preussen, xxii., p. 89, 
and Berg- und Huttenm. Ztg., 1872, p. 415. 



372 DESILVERIZATIOX OF BASE BULLIOX. 

i 136. Plattner's Cupelling Furnace.— In this connection may 
be mentioned the Plattner 1 modified German cupelling furnace. It 
has the form of a reverberators furnace; the hearth is rectangular 
in plan (13 by 8 J feet), and receives its blast from two pipes on 
either side of the fire-bridge ; the litharge-channel is at the oppo- 
site end, beneath the flue, which carries off the products of combus- 
tion as well as the lead-fumes. It would seem as if this modifica- 
tion might, with much advantage, be applied to English cupelling 
furnaces. Less fuel is required and less metal volatilized, because 
the litharge need not be heated to such a degree to remain liquid ; 
and no fumes enter the cupelling-room, as they are all carried off 
with the fuel-gases. 

§ is:. Mode of Conducting the Process. — The operations 

necessary to work a charge are six in number : preparing the 
working-bottom, charging &nd firing the furnace, softening the 
bullion, cupelling the softened bullion, removing the crude silver, 
and refining the crude silver. 

1. Preparing the Working-Bottom. — To be suitable for prepar- 
ing the hearth, the material must be one that will not be attacked 
by the litharge, nor crack, must be sufficiently porous to absorb 
some litharge, and free from any reducing agent (organic matter, 
metallic sulphide). The material most used is a marl. The com- 
position of that used in the Hartz Mountains varies according to 
Kerl and Wimuier 3 within the following limits. 

Si0 2 21.22 — 22.24 

A1 2 3 5.39 — 6.76 

Fe t O, 3.54 — 5.39 

CaO 65.65 — 66.41 

MgO 1.05 — 2.22 
Usually tae hearth-material is a mixture of dolomite or limestone 
with fire-clay. For instance, at Tarnowitz 3 a dolomite of the com- 
position, 



SiO a 6.00 




Al,O a 7.00 




Fe a O, 4.10 




CaOo, 49.86 




MgCo, 32.82 




'Drawings in Arche. '"Die Gewinnung der Metalle," Heft i., 


Plate i., 


Leipsic, 1888. Results in Berg- und Huttenm. Ztg., 1886, p. 211. 





1 Berg- und Huttenm. Ztg.. 1853, p. 241. 

s Zeitschrift fur Berg-, Butten- und Salinen-Wesen in Prussen, xxxii., p. 10" 



CUPELLAT10N. 373 

is mixed with 25 per cent, of clay. At Pribram three parts of 
limestone are ground together with one part of clay so as to pass a 5- 
mesh sieve; at other places, an 8-mesh sieve. The hearth -material 
has to be moistened before it is beaten down in the furnace. For 
this purpose it is spread on the floor, sprinkled with water from a 
rose, and turned over and over with a shovel, that the moisture 
may be equally disseminated through the powder. If left over 
night, it must be covered with wet cloths and worked again the 
next day. The material is of the right consistency if, when 
pressed in the hand, it coheres to a lump, but has not sufficient mois- 
ture to adhere to the hand. To obtain a uniform material, it is 
sifted, just before using, through a coarse hand-sieve, and any lumps 
that may have formed are broken up or thrown aside. Sometimes 
it is introduced all at once, sometimes in two separate layers. The 
latter is the way at Pribram, as Figure 255 shows, i being the 
upper and i' the lower bottom. Before the filling is put in, the 
brick bottom is sprinkled with water, that it may not take up any 
moisture from the hearth-material. At Pribram the lower bottom 
i' is first tamped down to the form shown in the drawing. The 
tool required, the tamping-iron, is a cast-iron disk of about six inches 
in diameter and one inch in thickness, with a socket into which fits 
a wooden handle about four feet long. The tamping is begun 
at the centre, proceeds in the form of a spiral to the side-walls, 
and returns in the same manner, care being taken that the circular 
indentations should overlap in part those made in working outward. 
By giving attention to this point, the surface will be evenly beaten 
down, which is essential. Before putting down the upper bottom, 
the surface of the lower one is roughened by scratching it with the 
point of a chisel. This is done that the bottom to be put down 
may adhere to the one already in place. The tamping of the upper 
bottom is done in the same way as the lower one, only the surface, 
when finished, must be perfectly smooth. It is of prime impor- 
tance that the hearth should have just the right degree of hardness. 
This is easily indicated to the ear and hand after a little experience. 
If too hard, it will crack and not be sufficiently porous ; if not 
hard enough it will absorb too much lead. If the material was 
too dry, the hearth will peel when heated ; if too wet, it cannot be 
beaten to the desired hardness. That it shall not adhere to the 
tamping-irons, these are slightly warmed. The thickness of the 
hearth at the bottom and sides varies somewhat ; the least is per- 
haps six inches at the bottom and eight inches at the sides. The 



374 DESILVERIZATION OF BASE BULLION. 

general rule for the curvature of the hearth is that the more concave 
the bed is, the easier will be the cupelling and the harder the finish- 
ing ; the flatter the bed, the harder the cupellation and the easier 
the finishing. When the hearth is completed, a cavity n (one inch 
deep and thirty-four inches in diameter) is cut in the lowest point 
to receive the silver. This is located a little to one side of the 
medial line, towards the fire-bridge, that the silver may be kept 
easily molten at the end of the operation. 

2. Charging and Firing the Furnace. — The furnace now 
receives its charge of 25 tons of lead enriched by Luce and Rozan's 
process. Sometimes the bottom is covered with straw before 
charging, to prevent its being damaged during the operation. The 
bars are placed in such a way as to leave an open space reaching 
from the tuyeres to the litharge channel. The hood is then 
lowered on a clay lute placed on top of the furnace. The litharge- 
channel is closed by lowering the door, the fire kindled on the 
grate, and soon the blast below let on. 

3. Softening the Bullion. — The lead melts down slowly. The 
dross rises to the surface and is drawn off through the litharge- 
channel. The temperature is raised and the blast put on through 
three tuyeres, the skimmings form and are drawn off, and finally 
pure litharge takes their place. 

4. Cupelling the Softened Bullion. — The temperature and the 
blast are now lowered, and are kept low during the larger part of 
the operation. They are raised only towards the end, when the 
enriched silver-lead alloy requires a higher temperature to give up 
the last parts of lead. "When the skimmings have been removed 
and the cupellation has somewhat progressed, the convex surface 
of the lead will be exposed to the action of the blast, while the 
lead near the periphery will be covered with litharge. The width 
of this rim depends on the rate at which the litharge is allowed to 
run off through the litharge-channel. 

As litharge melts only at 954° C. (§ 5), a temperature of about 
980° C. has to be maintained, if it is to remain liquid. Lead, melt- 
ing at 325° C. (§4), would be volatilized to a considerable extent, 
if fully exposed to the action of the blast at this high temperature. 
The litharge is therefore allowed to run off only to such an extent as 
to give the rim a width varying from 12 to 15 inches at the begin- 
ning, and of 5 inches toward the end, of the operation. The blast 
playing on the surface of the lead forms small waves and drives 
the litharge towards the channel. The pressure is about 8 ounces 



CUPELLATION. 375 

per square inch, and about 300 cubic feet of air are delivered per 
minute through the three tuyeres. In order to remove the litharge, 
a gutter is cut into the channel by means of a scraper. This is an 
iron rod, 8 feet long and f inches in diameter, flattened out at 
both ends. One of these is bent to encircle a wooden handle, 
while the other, only slightly flattened, is sharpened and bent to 
the form of a hook. In cutting the gutter, the entire edge of the 
tool must be used, and not one of the corners. If larger pieces of the 
breast are to be cut out, it is done with a chisel-pointed bar, say 
J of an inch in diameter. The rate at which the litharge runs off 
depends on the depth of the gutter and the strength of the blast. 
The depth is correct when the litharge runs off in a thin stream 
which stops as soon as the blast is lowered. If it runs too fast, 
the rim of litharge in the furnace decreases and lead is volatilized, 
while the litharge is not sufficiently desilverized from too short a 
contact with the lead beneath it. If it runs too slowly, the rim of 
litharge becomes too broad and the cupellation is retarded ; there is 
again loss in silver by the higher temperature that is required to 
keep the larger amount of litharge liquid, and if the temperature 
be not raised sufficiently, lead will be carried out mechanically by 
the litharge. The litharge gutter is first cut into the breast on the 
side farthest away from the bridge and gradually moved towards 
the opposite side, that the final litharge may be drawn off as near 
the fire-bridge as possible. The current of the litharge, when flow- 
ing out of the furnace, is directed in such a way as to form a large 
cake in front of the furnace. In some instances a U-shaped piece 
of sheet-iron is placed upright in front of the channel, that the 
litharge may collect in a rectangular block. Before removing this, 
the litharge in the centre that is still liquid is tapped from near 
the bottom of the cake. The litharge resulting from the cupel- 
lation is graded according to the silver-contents and the percent- 
age of impurities into marketable litharge and into a by-product 
to be treated by a separate process. As the cupellation progresses 
towards the end, the temperature is raised and the blast increased, 
the side tuyere-openings are closed, and two tuyere-pipes introduced 
through the central opening ; finally, the last film of litharge dis- 
appears from the surface of the lead with a characteristic phe- 
nomenon, the so-called brightening, 1 which every reader has 
watched while making a silver assay. 

1 Van Riemsdijk, Berg- und Euttenm. Ztg., 1880, pp. 247, 275, and Bock, 
Op. cit., 1880, p. 409, have made interesting investigations on this subject. 



376 



DES1LVER1ZAT10N OF BASE BULLION. 



The products of the Pribram cupelling furnace have, according 
to Dietrich, the following compositions : 





Dross. 


Tin Skim- 
mings. 


Antimony 
Skimmings. 


Red Lith- 
arge. 


Green 

Litharge. 


Cupel-Bot- 
tom. 


Flue 
Dust. 


Pb 


30.75 


13.40 












'PbO 


55 27 


64.97 


77.95 
11.87 


98.370 


98.140 


68.860 


64.41 


PbO com- 
bined with 
(AsSb) 2 5 . 














CuO 

Bi 2 3 

Mo0 4 

As 2 5 

Sb 2 5 

Sn0 2 

Ag. 

Ag 2 

A1 2 3 

Fe 2 3 

ZnO 

Ni 

NiO 

CaO 

CaC0 3 .... 

MgO 

C0 2 

Si0 2 

S '. 


1.99 

i'.ik 

1.83 
0.72 

'6!307 

I 0.54 

0.13 
0.09 

'6'.45 

Y.75 

2.30 
4.12 


0.29 

4187 

6.76 

10.31 

6 '.189 

[o.23 

0.05 

trace 

'6.43 

6.94 

0.65 
1.37 


0.28 

6192 

5.85 
trace 
0.004 

6.33 

6.14 

trace 

* 6164 

0.95 

"6!37 
0.09 
trace 


0.069 

'o'.oio 

0.074 

"61664 

0.072 

*6!6i6 

0.009 

'6.005 
0.256 

'61383 
0.320 

'6!634 


0.080 

'6 '.009 
0.067 

6. 0048 
0.056 

'6!6ii 

0.012 

oioos 

0.362 

"61432 
0.350 

'6I627 


0.070 

trace 
trace 

'6.530 

'6!i7o 

2.120 

"6!3o6 

trace 

24] 166 
trace 

'2.970 

61646 


trace 

| trace 
11.40 

616l3 

'6!so 

trace 

il35j 

4- ash j" 

16 1 65 

trace 


S0 3 

c 



Crude Silver. 





Pribram. • 


Freiberg. 3 


Wyandotte, Mich.' 


As 


95 
5 

trace 
trace 


92.180 
4.210 
2.104 
0.600 


98.691 
1.090 
0.117 
0.004 
0.090 
0.0058 
0.0023 


99.593 
0.260 
0.106 
0.008 
0.031 

'6.0015 


Pb 


Cu 


1 NiCo 


! Fe 


; Bi 


Au 





1 Loc. cit. 

3 Stolzel, " Metallurgies Brunswick, 1863-1886, p. 1182. 

3 Trans. A. I. M. E., ii., p. 97. 



CUPELLATION. 377 

The time required to cupel the twenty-five tons of rich lead is 
eighty hours. It is divided as follows : 

Hours. 
Preparing the hearth-material, making the hearth, and charging the 

lead 8 

Melting down and wheeling the necessary coal 16 

Drossing 6 

Drawing the tin skimmings 6 

Drawing the antimony skimmings 3 

Running off market-litharge 23 

Running off rich litharge 18 

80 

The cupellation is in charge of three men, each with a helper, 
working in eight-hour shifts. For every 100 tons of base bullion 
are consumed 19.63 tons of coal and 23 bushels of hearth-material 
(which includes the refining of the silver). The loss in silver is 
0.83 per cent.; that in lead, 4.33 per cent. The figures do not 
include the loss endured in re-treating some of the by-products. 
Thirty-six per cent, of the litharge is low enough in silver to be 
sold in the market, the bullion from which it is made averaging 167 
ounces silver to the ton. 

5. Removing the Crude Silver. — After the silver has bright- 
ened, the blast is shut off, the tuyere-pipes are removed, and the lith- 
arge gutter is closed with a ball of clay. Two knife-shaped pieces 
of wrought iron are introduced through the litharge-channel and 
pressed into the cake of crude silver. First warm, then cold, 
water is allowed to run into the furnace, and the silver then 
removed through the central opening at the back. It is cleaned, 
weighed, etc. The furnace is left to cool till the next day, when 
the hearth is examined for small particles of silver that adhered to 
it. The upper hearth is then removed with a pick. Part of it is 
soaked with litharge to a depth of 2 or 3 inches. This is screened 
off from the unsoaked part, which is mixed in with the hearth- 
material for the next charge, while the lead-soaked part goes to the 
blast furnace. 

6. Refining the Crude Silver. — The refining of the crude silver 
has for its object the removing of impurities, which vary from 2 to 
10 per cent. Formerly this was done exclusively in a small oval 
reverberatory furnace, having a working door at one side or at 
the flue-end, and a working bottom of similar composition to the 
cupelling bottom, the reason being that the loss in silver and 
the consumption of fuel were considered very much smaller than if 



378 DES1LVERIZATI0N OF BASE BULLION. 

the firing was done in the big cupelling furnace. Ohl ' and Foehr* 
have since proved this not to correspond to the facts, and the 
refining in a separate furnace has received a check. Since the dis- 
covery by Roessler 3 how to refine crude silver in a plumbago cru- 
cible by means of silver sulphate, the refining in a separate furnace 
has been abolished entirely at some works. 

Refining in a crucible is carried on at small works by melting 
down the silver, uncovering the crucible that the air may oxidize 
the impurities. These are stiffened by sprinkling bone-ash or 
hearth-material on the silver and then removed with a skimmer,, 
the operation being repeated till no more impurities rise to the 
surface. A slag obtained by Curtis 4 at Wyandotte, Mich., sand 
being used in refining, contained, in addition to silicate of lead, the 
following metals : 

(NiCo)O 0.550 



CuO 


0.203 


Bi 2 3 


0.026 


Ag 


1.837 


Sb 2 4 


0.639 


As 2 3 


0.005 



3.260 

Roessler found that if silver sulphate is added to melted silver 
in a crucible, first the lead and then the bismuth are converted to 
sulphates, the silver being at the same time set free. Copper is 
not removed by silver sulphate. By keeping separate the different 
slags he concentrates the bismuth in a comparatively small amount 
of slag, to be treated separately, while the first slag contains 
most of the lead. To prevent the crucible from being attacked, he 
introduces a layer of quartz-sand on top of the silver, and then 
stirs in the silver sulphate in the middle. The sand serves at the 
same time to stiffen the slag, which is then removed with a 
skimmer. The process, as seen by the writer at the Lautenthal 
Smelting and Refining Works in 1890, differs slightly from the 
manner indicated above. It is as follows. The silver sulphate is 
produced by dissolving silver in sulphuric acid of 66° B. in a small 
cast-iron kettle. The solution is allowed to cool, is then diluted to 

1 Berg- und Ruttenm. Ztg., 1879, p. 274. 

2 Ibid, 1885, p. 381. 

3 Berg- und Huttenm. Ztg., 1889, p. 387. 

4 Trans. A. I. M. E., ii., p. 98. 



CUPELLATION. 379 

60° B., when nearly all the silver sulphate will fall out as a slightly 
yellow cheesy mass. The supernatant liquor is drawn off as much 
as possible, and the remaining dilute acid driven off by heating. 
Special arrangements are required to cool the vapors, as they carry 
finely-divided silver sulphate along with them. The temperature 
is raised to redness in order to fuse the silver sulphate, which when 
liquid is cast into moulds and is ready for use. The color of the 
melted sulphate is grayish-green ; it is hygroscopic, and is therefore 
kept in a lead-lined wooden box ; 1,000 parts contain 650 parts of 
silver. 

Crude silver of a fineness varying from 950 to 980 thousandths 
is melted down in a plumbago crucible holding 700 pounds. The 
crucible is heated with coke in a small cylindrical furnace having 
in the lower part two lj-inch openings for the blast-pipes. On 
trying to stir in the silver sulphate as advised by Roessler, it was 
found that sometimes it got beneath the layer of sand, spread over 
the silver, and corroded the crucible. To prevent this, a wrought- 
iron ring (-J- inch thick, 10J inches in diameter, and 7 inches high) 
is coated on either side with a 3-inch layer of clay, and placed on 
the silver. Into the centre are introduced with a ladle from 6 to 8 
pounds of sulphate (the size of a hen's egg) that has been 
warmed. As soon as it comes in contact with the silver, this be- 
gins to boil. When the effect decreases, the silver is stirred with 
an iron rod to assist the action of the sulphate. From twenty-five 
to thirty minutes after the sulphate has been added, this is com- 
pletely decomposed, a slag has collected on the surface of the 
silver, and quartz is added to stiffen it, that it may be removed 
with a skimmer. A second, a third, and, if necessary, a fourth 
addition of silver sulphate is given to make the silver fine. The 
test made for fineness is to dissolve some silver into nitric acid and 
to supersaturate with ammonia. No precipitate must form even 
after standing. 

The amount of silver sulphate required to fine the silver is 
about 1£ times the total quantity base metal present. Thus 700 
pounds of crude silver, being 970 thousandths fine, contain 21 pounds 
of base metal, which would require 31 pounds of silver sulphate to 
be added in three portions. If the test with ammonia should 
prove this not to be sufficient, an extra addition is made. In 1890, 
107,031 pounds of crude silver with an average fineness of 970 
thousandths required 6,009 pounds of silver sulphate, which corre- 
sponds to about 2 parts of sulphate to 1 part of base metal. All the 



380 DES1LVERIZATION OF BASE BULLION. 

silver of the sulphate is not taken up by the silver in the crucible ; 
part of it enters the slag, as shown by an analysis made by Hampe. 1 



Si0 2 


40.7 


PA 


0.64 


so 3 


0.61 


s 


0.15 


FeO 


13.47 


A1 2 3 


0.43 


Bi 2 3 


6.01 


PbO 


33.50 


Ag 2 


2.05 ( = 1.88 Ag. 


Cu 


0.45 


Sb 


0.02 


CaO 


1.73 


MgO 


0.25 


K 2 


0.64 


Na,0 


0.26 



The main advantage of Roessler's method of refining is to be 
found in the larger direct output of silver that is obtained and the 
concentration of the bismuth in a comparatively small amount of 
slag that is more easily worked than cupel-bottom and litharge 
obtained in the reverberatory used for refining the silver. 

B. English Cupellation. 

§ 138. Characteristics. — The characteristics of this method are 
a small reverberatory furnace with a movable bed and a fixed roof, 
and the fact that the bullion to be cupelled is charged gradually 
and the silver is refined in the same furnace where the cupellation 
was carried on. 

§ 139. The Furnace. — This has undergone many changes from 
the original English furnace as described by Percy. 2 Figures 256- 
260 represent a form of cupelling furnace that is commonly used in 
American refining works. The vertical section (Figure 257) shows 
the general construction of the furnace, with the fire-place a, the 
vertical flue b, to the right, and the space c between the fire-bridge- 
wall d and the flue-wall e. This is closed at the top by the com- 
pass-ring f and the test when it has been put in place. The upper 
part of the furnace is encased in cast-iron plates ; the side-castings 

1 Berg- und Bftttenm. Ztg., 1891, p. 187 ; Engineering and Mining Journal, 
November 14, 1891. 

2 "Metallurgy of Lead," p. 178. 



CUPELLATION. 



381 



have a strengthening rib to resist the thrust of the roof. In 
addition, the front g of the ash-pit A, the flue-end of the furnace, 
as well as the inner sides of i andy, are protected by castings. 
The usual buckstays and tie-rods have been left out in the draw- 
ings. To be noted are the large grate area (4 feet 6 inches by 2 



Fig. 256. 



Fig..25.7\ 



ya 


" r "~ r -"*- : ;;7-----^-i4'J 


■**& 'fl 


1 II 


30 3 

: e 

>■ 


i- 


-V 

10-7- 




Fig. 358. 




SECTION ON C D, 
7 1" k' P i 7 5 




■fr-fc 



SECTION ON A B. 
Fig. 359. 



feet J inch) in comparison with the hearth area (4 feet 4 inches by 
3 feet 6 inches), the height from the grate bearer k to the roof I (2 
feet 4 inches) and the short distance (9-J- inches) between the roof 
and the top of the compass-ring, as they are all essential for good 
working. The flame from the grate-bars is directed downward by 
the pitch of the roof, and, being forced to pass through the small 

Fig.. 2 6 O. 

SECTION ON E F. 












KlDI — 11 Tu 


■>~ 








*^ 



space between the shallow roof and the hearth, exerts all its heat- 
ing force on the lead. By paying special attention to this part of 
the construction, it is possible to obtain a sufficiently high tempera- 
ture to refine silver without being forced to use special kinds of 



382 DESILVERIZAT10N OF BASE BULLION. 

bituminous coal. In fact, with forced blast under the grate, slack 
coal is good enough for cupelling, and nut coal is required only for 
fining, the coal being of ordinary grade. The grate in the draw- 
ing is so arranged that firing, as well as stoking, is done from the 
front. Another plan is to place the grate-bars parallel with the 
short sides of the fire-place, and to stoke from the short side of 
the furnace. If in such a case forced blast is used, the casting at 
the stoking side has two oblong openings, each to be closed by a 
cast-iron door, say 21 inches long and 6 inches high, having its two 
hinges on the lower side. The horizontal flue m ( Figure 257) is 
seen in Figures 259 and 260 to be divided into three smaller 
flues m', increasing in width (11, 13, 16 inches) from the back of 
the furnace toward the front, the object being to prevent the flame 
from taking the shortest line toward the centre of the flue, and to 
draw it somewhat toward the front, whereby the litharge floating 
on the front part of the lead is kept hot. At some works this flue 
is divided into five or six smaller flues. If in the furnace de- 
scribed the conditions of the draught are such that the flame rushes 
too much toward the centre, this is remedied by placing one or 
more fire-bricks in the flue, which will correct the evil. At the 
back of the furnace thei*e is only one door n, through which 
passes the blast-pipe and through which is fed one pig of lead at a 
time. Many furnaces have three small openings, a central one for 
the blast-pipe and two lateral ones, through each of which a pig of 
lead is gradually pushed forward and melted away. Having only 
one opening, to be closed by a sliding door running in a vertical 
frame o, simplifies the casting. The compass-ring, whose inner 
contour must, of course, correspond to the form of the test that is 
to pass through it, is intended for a rectangular cast-iron test 
having rounded corners. The upper rim p serves to hold in 
place the brickwork with which its surface is covered. It ex- 
tends all around the central opening, with the exception of the 
front, where it stops for a distance of 16 inches, leaving room for 
the slot g (4 inches wide), through which the litharge is to run 
down into the litharge-pot. The outer contour of the compass- 
ring has the rectangular form of the furnace, and reaches from 
front to back. It thus forms at the front the support for the 
working tools, and at the back the support for the prop which 
gives the blast-pipe the desired pitch. At the front the compass- 
ring is often left open the entire length of the litharge-slot. In 
such a case the support for the tools is a special cross-bar (skim- 



CUPELLATION. 



383 



ming-bar) held in place by screws, or by being let into the brick- 
work, or fastened in some other convenient manner. 

§ 140. Test-Rings. — The test (4 feet by 2 feet 6 inches), which 
originally consisted of an oval frame of wrought-iron (4J- inches 



Fig. 263. 

SECTION ON C D. 
51- 




wide, J inch thick) filled with bone-ash, has undergone many 
changes in construction, manner of support, and filling material. 
A few of the tests in use at present are represented in Figures 261- 
269. Figures 261 and 262 show a cast-iron test i resting on a test- 



Fig, 263. 

BOTTOM DOWNWARDS 



Fig. 264. 
BOTTOM UPWARDS 




SECTION ON A B, 

Fig. 364a. 




si^ 



Fig. 266 






( l"pipe Inlet 



3f4 Pipe Outlet 
1 



carriage. It has a concave bottom and a cast-iron pattern of the 
cavity on the inside, both of which will be discussed farther on. 
The test-ring is rectangular in plan and has rounded corners. It 
fits into the cupelling furnace shown in Figures 256-260. When in 



384 DESILVERIZATION OF BASE BULLIOX. 

place, its upper surface will be on a line with the upper rim of the 
compass-ring p (Figure 257). In front it has a 3-inch slot J which 
is closed when the filling material forming the hearth is being 
rammed in. The rectangular form of test offers a large surface 
for oxidation, and therefore more lead can be cupelled on it than 
on an oval hearth. 

Figures 263-264« represent an oval cast-iron test-ring, having a 
horizontal flange a. When in place, the upper side of this will be 

Tig.267. 
HoopA< r Iron Pipe 1* 

Iron Bind I 



'J7$ 



Fig.268. 



uiin2-- ^7> 

/ PLAN. 

4M 1^1 ELEVATION. 

Casting >» thick 
FiG.269. 

close to or in contact with the lower side of the compass-ring f 
(Figure 257), and the upper edge of the test-ring itself will be in 
line with the upper edge of the compass-ring, thus making the 
distance between the roof and the surface of the lead as small as 
possible. The test-ring protrudes over the horizontal flange at c, 
forming a loop d. In tamping down the filling material the loop 
is not filled, in order that the litharge overflowing from the hearth 
may pass through it into the litharge-pot below ; thus any contact 
between hot litharge and iron frame is avoided. Across the bot- 
tom of the test-rinor are four cast-iron arms e to hold the filling 
in place. 

With both tests the corrosive action of the litharge very soon 
eats out the filling, especially at the front, and in a compara- 
tively short time the test has to be removed from the furnace 
and replaced by another. To counteract the corrosion of the 
front, and at the same time permit the raising or lowering of 
the litharge-gutter, a water jacket, as shown in Figures 268 and 
269, is fastened by means of bolts to the test-ring, be this cast 
iron or wrought iron. The jacket has the same depth as the test- 
ring and forms the breast. The litharge runs off through gutters 
cut into the filling b (Figure 269). As this is cooled by the water 
circulating in the jacket, it is eaten out only very slowly. The 



CUPELLATION. 385 

jacket itself does not come in contact with the hot litharge in 
the furnace, as it is protected from it by a 3-inch rim of filling. 
This wears out somewhat, but never or rarely so far as to bring 
the casting into direct contact with the hot litharge. 

By this arrangement only the front is protected. A device 
that protects the sides alone is shown in Figure 267. Here a is a 
cast-iron test-ring, resting on a bed-plate b. It is surrounded by 
an iron hoop tied by an iron band. On top of the test-ring are 
placed two 1-inch pipes in which water circulates. The filling d 
is rammed down in the usual way. The lead, while the furnace is 
running, being always kept about at the same height, the litharge 
can show its bad effect only on the level of the water-pipes, and 
these effectively counteract to a great extent the corrosive action. 

A combination of the arrangement shown in Figure 267 with 
those in Figures 268 and 269 might protect both sides and front of 
the test-ring. 

This is effectively done by the Steitz water-jacket test, as repre- 
sented in Figures 265 and 266. Here a is a rectangular water jacket 
made of boiler-iron. The open space in front is closed by a cast- 
iron water jacket having a litharge-gutter e ; it is fastened with bolts 
to the wrought-iron jacket. The jackets are placed on a cast- 
iron test-plate c, which supports the filling d. Here both jackets 
are protected from the hot litharge forming on the surface of the 
lead in the furnace. The gutter e alone is attacked by the lith- 
arge, and is eaten out after some time; the breast jacket then has 
to be exchanged for another. This is quickly done, while the 
furnace is being used, without any difficulty whatever. The effect 
of jacketing is that the filling lasts longer than with the other 
test-rings. The cast-iron front jacket, in addition to preserving 
the breast better than any of the other arrangements, has for 
concentrating another advantage, that the depth of the litharge- 
gutter does not have to be regulated by the cupeller, but always 
remains the same. This renders it, however, unavailable for bring- 
ing very rich, say 70 per cent., bullion up to fine silver, as the uni- 
form level of the gutter prolongs indefinitely the removal of the 
last lead contained in the silver. 

§ 141. Test-Ring Supports. — The manner of bringing the tests 
into position and fixing them there has of late years undergone 
various changes. The old method consisted in driving four iron 
wedges between the bottom of the test-ring and two transverse 
bars, the ends of which were inserted 4 inches below the test- 
25 



386 



DESILVERIZATION OF BASE BULLION. 



frame into the fire-bridge wall d and the flue wall e (Figure 258). 
It is not much used now. 

Sometimes wedges are still retained to adjust a large test-frame, 
as represented in Figure 265. This is placed on two 9-inch brick 
walls running along the fire-bridge wall and flue wall. It is then 
raised gradually to the height of 12 inches, and four pillars, each 
three bricks high, are placed beneath the bed-plate, which brings it 
up nearly to the compass-ring. By then driving flat wedges 
between the bed-plate and the brick posts the test is adjusted to its 
final position. A common method of supporting the test is by 
four screws (18 inches long, 1J inches in diameter), working in 
two transverse bars placed 12 inches beneath the test-ring. This 
rests on a cast-iron plate into which the points of the four screws 
are set. 

Four screws are also found in connection with a test-carriage, as 
shown in Figures 270-272. x Here the test is easily brought into 



Fig. 271. 




position, and then raised by means of the screws and brought up 
close against the compass-ring. 

With the three arrangements described, the test, when once 
placed, is immovably fixed while the cupellation is proceeding. 

1 Taken from a drawing of Messrs. Fraser & Chalmers. 



CUPELLATION. 



38? 



Further, when the furnace has become hot, the turning of the 
four screws often presents considerable difficulties. To obviate 
these disadvantages the two front screws have been removed and 
the test suspended or held by a support. 

The Lynch ' test-support is represented by Figures 273 and 274. 




]Fig. 274. 



a Pulley Block 

6 Turn-Bolts 

c Support-Bars 

d Back Screws 

e Test Plate 



The test plate e, shown bottom side up in Figure 274, has at the 
back the two sockets for the points of the two back screws d 
(Figure 273). Two support-bars c, diverging 27 inches when 
extending in front of the furnace, are clamped to the front half of 
the plate. When in place, they are hooked with the turn -bolts b 
to a triangle made of f -inch iron, which is attached to a differential 
pulley a hung from the roof. With the pulley the test can be 
raised or lowered quickly and evenly to regulate the flow of 
litharge without altering the depth of the litharge-gutter. The 
contents of the test (rich lead or silver) can also be poured. With 
the turn-bolts it can be moved sideways to counteract the action of 
the litharge, should this corrode one side of the hearth more than 
the other. 

1 Blake, Trans. A. I. M. E., x., p. 220. Patent No. 275,232, April 3, 1883. 



388 



DES1LVER1ZATI0N OF BASE BULLION 



Another movable test-support is represented by Figures 261, 
262, 275. The test is supported by a carriage o, with its movable 
upper frame act ' . This rests at the back on two screws e and e', 
working in the blocks d and cl. At the front, it is supported by 
the screw g, working in the right arm a of the frame, which 
extends a short distance in front of the furnace, and is turned out- 
ward sufficiently for the wheel h to be to the right of the cupeller. 
By inserting a hook into one of its circular openings he turns the 
wheel to right or left, and thus raises or lowers the front of the 




SECTION ON Q H. 



frame aa' and with it that of the test. Into the upper frame cross- 
beams b b' b", riveted to each other in pairs, are let in, which 
serve as support for the concave bottom of the test. If this were 
straight the surface of beam a would represent in the section an 
unbroken line, two more beams like c and c' being sufficient to give 
the frame the required stability. The wheels of the carriage are 
grooved. By having two rails running across the space c (Figure 
256), all shifting of the carriage to right or left to get it into correct 
position is avoided. 

The support at the back by the screws e and e has the disad- 
vantage that when the frame aa' has to be lowered there before the 
carriage is taken out, much difficulty is experienced in turning the 
screws. To avoid this the screws e and e have in one instance 
been replaced by two pivots which, being fastened into the blocks 
d and d', fit into circular openings of two blocks fastened to the 
beams a and a'. When the frame, resting at the back on these 
two pivots and supporting the test, has been w T heeled into its cor- 
rect position and is to be raised, this is done by lifting it a few inches 
with a crowbar ; then two U-shaped castings, of the correct height, 
and wide enough to enclose a pivot, are placed around these. On 
withdrawing the crowbar the frame sinks on these two castings, 
which then support it. When the test is to be exchanged, the 



CUPELLATION. 389 

frame is again slightly raised, the two U-shaped castings are 
thrown off, and the frame is slowly lowered, when it will again 
be supported by the pivots. The test-ring is then sufficiently- 
low for the carriage to be withdrawn from under the compass- 
ring. 

Another movable support with specially constructed test-ring 
has been invented by Roesing, 1 and is used at Tarnowitz, Silesia. 
Movable tests are in much favor when the test-ring is not water- 
cooled, i.e., where the level of the lead is gradually lowered by cut- 
ting deeper the litharge-gutter ; with concentrating tests that are 
entirely jacketed they are not used, and with those that are water- 
cooled at the breast the regulating of the litharge-gutter has not 
caused such difficulties as require a movable test. 

Whatever test-support may be in use, care is always taken to 
plaster over with clay the upper surface of the ring, so as to pro- 
tect it against coming in direct contact with the flame, and at the 
same time to prevent the air from rushing in between compass-ring 
and test-ring. 

§ 142. The Blast. — The blast was originally produced by a 
steam-jet. This has given way entirely to a blower. The ma- 
chines in common use are the Baker and Root blowers, and the 
Sturtevant fan ; blower and engine are usually supported on the 
same bed-plate. The pressure of the blast is about four ounces 
per square inch. The blast-pipe is usually of sheet iron and is 3 
inches in diameter. At some works it fits into a cast-iron nozzle, 
which delivers the air through an aperture 4 inches long and \ inch 
wide. This is better than simply flattening the end of the sheet- 
iron pipe. 

§ 143. The Tools. — The tools required by the cupeller are few : 
2 rods (7 feet long, of J-inch iron), bent to a hook ; 1 chisel- 
pointed bar (7 feet long, of J-inch iron); 1 saw (9 inches long, 
^ inch wide, 2 inches deep), attached to a handle (6 feet long, of ^~ 
inch iron), to cut down the breast ; 1 fire-hook (10 feet long, of 1- 
inch iron, with a 4 by 10-inch head); 1 scoop ; 1 ladle ; 10 bullion- 
moulds or silver-moulds, and 3 litharge-buggies having small ket- 
tles (13 inches in diameter and 8 inches deep) to hold the litharge. 

§ 144. Mode of Conducting the Process.— The operations in- 
clude filling the test and putting it in place, cupelling, and refin- 
ing. 

: Berg- und Huttenm. Ztg,, 1883, p. 577 ; Engineering and Mining Journal, 
January 19, 1884. 



390 DESILVER1ZATION OF BASE BULLION. 

1. Filling the Test and Putting it in Place. — The material 
originally used to fill the test was bone-ash ground fine enough to 
pass a 26-raesh sieve. This, being too expensive, has given way to 
a mixture of limestone and fire-clay ground through a 12-mesh 
sieve, the proportions being three parts by volume of limestone to 
two or sometimes only one part of clay, according to the plasticity. 
Portland cement is used at some works, and if of best quality is 
more durable than the limestone-clay filling. Sometimes a mixture 
of two-thirds cement and one-third coarsely ground fire-brick is 
used instead of pure cement. The test has also been brick-lined ; 
experiments, which are very promising, are going on with magnesia- 
brick. 

In filling the test-ring with the limestone-clay mixture, this is 
moistened, as shown in § 134, and then tamped into the ring. 
Some works beat down the mass in three separate layers ; others 
add the necessary material all at once and begin then with the 
tamping. If the cast-iron test-ring (Figures 263-264«) is to be 
filled, a piece of wood having the form of the loop d is put in 
place, and then the filling beaten down. When finished, the wood 
is taken out, leaving open the slot for the discharge of the litharge. 
The wood is best withdrawn only when the test has somewhat 
dried, as then there is less danger of breaking off part of the filling. 
The tamping-irons are about 4 inches in diameter at the base. The 
test is filled entirely with the material, and the cavity then scooped 
out with a trowel. A very good way to insure a uniformly hard 
surface is to place a wooden frame on top of the test-ring, and 
then partly fill it with the hearth-material. When the frame has 
been removed, the excess of material is scraped off down to the test- 
ring, and the cavity then scooped out. A rim from 3 to 4 inches 
wide at back and sides, sometimes (5 and even 10 inches at the 
front, is left untouched. The cavity has its lowest point near the 
front to facilitate the dipping out of concentrated bullion or fine 
silver. The filling material should be at least 4 inches thick at 
the lowest point, and the depth of the cavity about 5 inches. 
Thus an oval test 4 feet 6 inches long, 2 feet 1 inch wide, and 5 
inches deep, holds about 2,500 pounds of lead. • 

In filling the test with Portland cement or with a mixture of 
cement and ground brick, this is moistened and tamped down in 
the usual way, the cavity, however, being formed during the tamp- 
ing. A quicker and better method is to place the test bottom 
upwards over a mould having the form of the cavity, and then to 



CUPELLATION. 391 

beat down the moistened cement. In using cement it is essential 
that the work be done quickly, as it must be finished before the 
cement shows any signs of setting. 

The test frame i (Figures 261 and 262) is the only one that has 
a solid cast-iron bottom ; it is slightly concave. The working 
bottom is made of fire-brick, which is set dry and then thoroughly 
grouted with fire-clay. To build up the sides, the cast-iron frame 
k is oiled and put in place, and moistened cement is beaten down 
in the intervening space v, to the top of the test-frame ; the iron 
frame is then carefully removed. It is in this test that an entire 
lining of magnesia-brick promises good results. A test of the 
same description is used to concentrate copper matte, obtained from 
melting copper dross with galena (§ 119), into bottoms, and 60 per 
cent, matte, which is then converted into metallic copper. 

When a test is filled, it has to stand for a fortnight and longer 
in a warm place (usually the cupelling-room) to dry. Before it is 
to be used in a cold furnace, a small charcoal fire is made on it. 
In a warm furnace the fire is kept low for three or four hours after 
the test is in place. 

2. Cupelling and Refilling. — When the test is in place and 
well warmed, the temperature of the furnace is gradually brought 
to a dark-red, and some lead introduced through the front and 
melted down. When this has become a cherry-red, the blast is let 
on and cupellation started. No distinction is made between dross, 
skimmings, and litharge, as in the German cupellation. The 
litharge is made to run off at the front, and fresh lead is supplied 
from the back, where one or two small bars protruding through 
openings into the furnace are melted down at such a rate as to 
keep the lead in the test always on the same level. The litharge 
is collected in a cast-iron pot (say 13 inches in diameter and 
8 inches deep) running on wheels. The litharge-pot was at one 
time replaced by a water-box. This has two advantages. It 
reduces the temperature for the cupeller and presents the litharge 
in a granulated form, which is easily handled and sampled. It 
has, however, the disadvantage that a cupel-carriage cannot be 
used, and that in the blast furnace there is more loss in lead and 
silver when granular litharge is charged than if it is in lump form. 
Granulating has been abandoned. 

With a stationary iron test-ring, the litharge is run off through 
a gutter cut into the filling. It is not often that one gutter serves 
for the passage of the litharge ; generally there are three and. 



392 DESILVERIZATION OF BASE BULLION. 

often four, opened one after the other to prevent excessive corro- 
sion. A movable test gives an additional way of regulating the 
flow of litharge by lowering and raising the front. With the 
Steitz water-jacket test the flow of the litharge is regulated only 
by the quantity of the lead that is melted off from the bars at the 
back of the furnace. The gutter can be closed for a short time by 
allowing litharge to accumulate there, or with a piece of clay. 

The flow of the litharge is so regulated that about one-half 
the surface of the lead remains covered. The former practice of 
cupelling and fining in the same furnace has been abandoned at 
most works. At present it is common to concentrate the bul- 
lion to 60 or 70 per cent, of silver on one test and to fine it in a 
separate furnace. For this concentrating, the water-jacket test is 
excellent, as it can be run by an inexperienced man, while judgment 
and practice are necessary with a test where the litharge-gutter 
has to be regulated by the cupeller. By thus dividing the cupel- 
ling into the two operations of concentrating and finishing, a smaller 
number of experienced and reliable cupellers is necessary. When 
the bullion is concentrated to the desired degree, it is ladled out 
and goes to the finishing furnace, and the concentration furnace is 
again filled. Thus a concentrating furnace runs constantly. After 
a certain time the bottom becomes too thin and has to be exchanged. 
A test-ring filled with limestone-clay, if used four or five hours 
daily for finishing, lasts only thirty days ; a cement-test used for 
the same purpose lasts months. A water-jacket test filled with 
limestone-clay, used for concentrating, lasts only sixty days. A 
test-ring filled with cement and used for concentrating and refining 
lasts seven days. 

The finishing is always done on a test having an iron test-ring. 
The operation is the same as in concentrating, but usually not con- 
tinuous. Towards the end, when the silver-lead alloy becomes 
less readily fusible, the temperature has to be considerably raised. 
When the silver has become sufficiently concentrated, the addition 
of rich bullion is stopped. The last litharges are drawn off and the 
test remains almost filled with crude silver, which has now to be 
fined. It is not often that the brightening is seen. Samples taken 
from the metal bath show how far the cupellation has progressed. 
The fining consists usually in exposing the silver for some time to 
the action of the heat and the blast. Bone-ash is sometimes given 
in small quantities to absorb the impurities that float on the surface 
or have collected on the edge. At some works nitrate of soda 



CUPELLATION. 393 

is used in the furnace to make the silver at least 997 thousandths 
fine, as this has become the standard below which fine silver should 
not go. The nitre is spread on the silver, a shovelful (about twelve 
pounds) at a time. To prevent the soda from corroding the filling 
of the test-ring, some refiners spread finely ground brick over the 
nitre. The slag, floating on the silver, is removed only when the 
silver is fine and ready to be cast into moulds. About 15 shovel- 
fuls of nitre are required for 50,000 ounces of silver. 

The indications of fine silver are : a smooth, clean surface ; 
that stirring fails to bring impurities to the surface ; that a 
tool held over the silver is clearly reflected in it ; that a sample 
taken by dipping in a rod will show no spots whatever on the sur- 
face, and have a pure, silver-white color ; and that a sample taken 
with a spoon will spurt while cooling, although this is not a 
good test. Some refiners cast a small sample-bar, examine the 
surface, which should be smooth ; the fracture, which should be 
finely granular and show a silky lustre ; test for malleability by 
hammering, etc. The only way to know definitely how the fining 
is progressing is to make an assay. This is done in the dry way, 
by weighing out twice J gramme of a granulated sample and £ 
gramme of c. p. silver as a check, and cupelling the three samples 
with the same amount of lead on three cupels placed in a row in 
the muffle. A second assay half an hour later will show whether 
any progress has been made. An assay in the wet way, with 
potassium-sulphocyanide, using ferric sulphate as an indicator, 
will give the same result quicker than cupelling. 

When the silver is fine, that is, when it ranges between 997 and 
999.5 thousandths, it is either ladled out into warmed moulds, or, if 
the Lynch test-support is used, it can be poured. Sometimes this 
is done into water, to be remelted at a lower temperature in a 
plumbago crucible or a new retort (heated in a tilting furnace), and 
then cast into moulds. This gives smoother bars than would be the 
case if ladled or poured from the test, but otherwise has no special 
advantage. Smooth bars are, however, demanded at present. 

3. Sampling and Assaying Fine Silver. — The sample of the 
fine silver is best taken from the mould. When this has been 
filled, a long-handled iron spoon is inserted, the silver stirred with 
it, and the sample taken out and granulated in water. Taking 
chips from different parts of the bar is unsatisfactory, as the 
impurities are as a rule not evenly distributed. 1 In assaying the 
1 Blake, Journal of Chemistry, 1888, p. 71. 



394 DESILVERIZATION OF BASE BULLION. 

fine silver, either the Gay-Lussac method or the Vollhard method 
is used. The latter gives with the necessary care very satis- 
factory results, notwithstanding adverse criticism. 1 Sometimes 
the dry assay alone, as indicated above, is made to serve as a 
check for the mint returns. Details of these wet methods are 
given in every book on assaying. A steam bath used for dissolving 
the silver and an apparatus for shaking the assay bottles, have 
been fully described by Blake. 3 In assaying dore bullion for gold, 
it is advisable to weigh out and dissolve, without previous cupel- 
ling, a separate sample. The amount of gold present rarely exceeds 
14 thousandths. 

4. By-Products. — The by-products of the cupellation process 
are litharge, cupel-bottom, and flue-dust. Litharge coming from 
retort bullion runs, when it is pure, from 50 to 60 ounces to the ton ; 
when it is impure, e.g., when drosses of the retort bullion are 
being scorified, often from 150 to 200 ounces. The cupel-bottom 
varies too much in lead and silver to give any average figure. 
A sample of flue-dust assayed 21.6 per cent, lead, 20 ounces silver, 
and 0.16 ounces gold. These by-products always go to the ore 
blast-furnace ; litharge is sometimes used, as already stated, to 
hasten the softening of base bullion that is especially hard. 

5. Results. — In a test 4 feet 6 inches by 3 feet 6 inches, 
7,000 pounds of retort bullion are cupelled by two men in twenty- 
four hours, using from \\ to 2 tons of coal, according to the 
quality of the fuel. It is advisable to have 8-hour shifts for 
cupellers to prevent their becoming leaded. 

The concentrating of 1,000 pounds of 70 per cent, bullion on a 
33 by 28-inch test, 5 inches deep, and refining of the resulting 
silver (say 12,000 ounces silver), lasts about five hours, requires 
one man, and about 1,500 pounds of nut coal. When the finish- 
ing furnace is stopped for a day or two, the fire on the grate is 
kept going in order that the temperature of the test may not sink 
below a dull red heat ; charcoal is often kept aglow on the test, as 
it makes it again porous when it is much soaked with litharge. 

The loss of lead in cupelling is generally given as 5 per cent. 

§ 145. Comparison of Methods. — A comparison between the 
two methods of cupelling leads to the conclusion that the German 
method is, for purposes for which cupelling is generally used to- 
day, by far the more expensive. Although it forms litharge more 

1 Torrey, Engineering and Mining Journal, January 6, 1883. 

2 Trans. A. I. M. E., x., p. 492. 



CUPELLATION. 395 

rapidly, because the hearth is so much larger, it produces only a 
comparatively small amount of silver as the product of one opera- 
tion. To remove the silver, the furnace has to be cooled, and the 
hearth torn out and replaced by a new one. This takes time, thus 
neutralizing the advantage of the quick formation of litharge, and 
costs much labor, fuel, and material, including a large amount of 
hearth-material, which has to be smelted in the blast-furnace for 
every cupellation. In the English cupelling furnace, especially 
with its American modifications, a cupel-bottom can be made to 
last for weeks ; the process is therefore less interrupted, and thus 
much expense for labor, fuel, and material saved, but it has the 
drawback that the litharge is always more apt to be rich and 
impure. 

A German cupelling furnace may be in place when the result- 
ing litharge is to be sold as such, and is therefore required to be 
pure, and poor in silver. The English furnace with American 
improvements is decidedly preferable if the bullion to be cupelled 
is so rich that the resulting litharge would in any case run too 
high in silver to be sold as such. In this case, and it is the com- 
mon one to-day, it is not of much consequence whether the litharge 
be a little poorer or richer in silver, or if it be somewhat con- 
taminated with impurities, as long as the advantages more than 
make up for such deficiencies. 



INDEX. 



{Figures refer to pages, not to sections.) 



Accretions : 

In the hearth, 280. 
On furnace-walls, 278. 
Agglomerating of fine ore, 157, 293. 
Air-Furnace: 

Analyses of products of the, 90. 
Lead-smelting in the, 89. 
Air-reduction process, the, 20, 79. 
Aitken on hot plates repelling dust 

particles, 286. 
Alder-Wright on alloys of zinc and 

silver, 310. 
Alloys : 

Of lead, 21. 
Of zinc and silver, 310. 
Altitude, effect of, on the amount of 
fuel required in the blast-fur- 
nace, 154. 
Alumina : 
Action of, in the blast-furnace, 143. 
Method of analysis, 277. 
Aluminum, use of, in zinc-desilveriza- 

tion, 329. 
American ore-hearth, the, 114. 
American water-back ore-hearth, the, 

113. 
Analyses of: 

Antimony-skimmings from cupella- 

tion, 376. 
Base bullion, 244. 

Black slag from the slag-eye fur- 
nace, 124. 
Blast-furnace gases, 207. 
Blue-powder from the ore-hearth, 128. 
Carbonate lead-ore: 
From Leadville, Col., 38. 
From the Hornsilver mine, Utah, 42. 
Old Telegraph mine, Utah, 42. 
Red Mountain district, Col., 33. 
Richmond mine, Xev., 41. 
Coke-ash, 151, 210. 
Crude silver, 376. 
Cupel-bottom, 376. 
Dross from hard lead, 361. 
From softening, raw, and melted, 
315. 



Analyses of : 

Dross from the cupelling-furnace,. 

376. 
Flue-dust from the blast-furnace, 
281. 
From the cupelling-furnace, 376. 
From the ore-hearth. 128. 
From the reverberatory-furnace, 

Tarnowitz, 105. 
From the roasting-furnace, 167. 
Hard lead, 361. 
Hearth accretions, 280. 
Hearth material for the cupelling- 
furnace, 379. 
Iron fluxes, 139, 210. 
Lead before and after drossing, 
315. 
Of commerce, 21, 87, 90, 118, 367. 
Lime fluxes. 142, 210. 
Litharge, 376. 
Manganese fluxes, 140. 
Matte, 253. 
Matte and matte-slag from Clausthal 

and Freiberg, 263, 265, 266. 
Matte, raw and heap-roasted, 256. 
Matte, raw and stall-roasted. 260. 
Oxides, from steaming zinc-bearing- 
lead, 341. 
Residue from the reverberatory-fur- 
nace : 
Engis, 89. 
Missouri, 90. 
Tarnowitz, 105. 
Roasted blue-powder from the ore- 
hearth, 118. 
Roasted galena ore, 167. 
Silver-refining slag, 378, 380. 
Slag from the ore-hearth, 118. 
Sows, 280. 
Speise, 250. 
Sulphide lead-ore : 
From Engis, Belgium, 87. 
From Leadville, Col., 37. 
From Mine La Motte, Mo., nickeL 

bearing concentrates, 32. 
From Raibl, Carinthia, 84. 
From the Aspen mines, Col., 39* 
From Caribou mine, Col., 33. 



398 



INDEX. 



Analyses of : 

Sulphide lead-ore : 

From the Colonel Sellers mine, Col., 
raw ore and concentrates, 38. 

Tin skimmings, 361, 376. 

Wall accretions, 279. 

White paint from the slag-eye fur- 
nace, 130. 

Zinc crusts, 327. 
Anglesite, 27. 
Antimony : 

Action 'of, in the blast-furnace, 149. 

Effect of, on lead, 23. 

In cupellation, 369. 

Influence of, in the roasting and re- 
action process, 83. 

In zinc-desilverization, 312. 
Antimony-skimmings from the soften- 
ing furnace, 323. 
Antimony-skimmings, smelting of, in 

the blast-furnace, 359. 
Argillaceous matter, influence of, in the 
roasting and reaction process, 82. 
Arents' automatic tap, 187, 189. 
Arizona, lead ores of, 44. 
Arsenic : 

Action of, in the blast-furnace, 149. 

Effect of, on lead, 23. 

In cupelling, 369. 

Influence in the roasting and reaction 
process, 83. 

In zinc-desilverization, 312. 

Method of assaying, 252. 
Arsenical skimmings from the soften- 
ing furnace, 323. 
Ash in cote, analysis of, 151, 210. 
Ash in coke, per cent, of, 150. 
Assay: See also Determination. 

Book, form of, 243. 

Calculation of average of, 64. 

Crucibles, 70. 

Furnaces, 69. 

Method for copper, 271. 

Method for gold, 72. 

Method for lead in matte and base 
ores, 270. 

Method for lead in ores, 69, 270. 

Method for silver in base bullion, 248. 

Method for silver in ores and slags, 
69. 

Method for silver in silver bullion, 
393. 

Sample of ores, 47. 
Associated minerals of galena, 26. 
Atlantic coast, lead ores of the, 30. 
Atmosphere, action of the, on lead, 15. 
Austin on sampling base bullion, 247. 
Automatic sampling machines, 52. 
Automatic tap, 187, 189. 
Available iron for fluxing, 211. 
Average assay, calculation of, 64. 



Bag-process, the Lewis and Bartlett, 

124. 
Baker on the effect of silver on lead of 

commerce, 22. 
Balling, slag-table of, 134. 
Bauer on the effect of bismuth on lead 

of commerce, 22. 
Balbach process, the, 348. 
Barite: 

Action of, in the blast-furnace, 114. 

Influence of, in the roasting and re- 
action process, 82. 
Barring down of wall accretions, 236. 
Bartlett. separating zinc from lead ore, 

148. 
Baryta : 

Manner of figuring in a blast-fur- 
nace charge, 145. 211. 

Method of analysis, 275, 277. 
Base bullion: 

Analyses of, 244. 

Bessemerizing of, 296. 

Desilverization of, 296. 

Receiving of, 314. 

Requirements of, for zinc-desilveriza- 
tion, 311. 

Sampling of, 246. 

Softening of, 314, 322. 
Base excess, definition of, 73. 
Basic lining in cupel tests, 390. 
Batopilas, channelling at, 52. 
Battersea Company, crucibles of the, 70. 
Battery assay for copper, 272. 
Berthier on alloys of nickel, cobalt, and 

lead, 23. 
Bessemerizing base bullion, 296. 
Binon and Grandfils, separating of zinc 

from lead ore, 147. 
Bins and beds, records of, 68. 
Bismuth : 

Effect of, on lead of commerce, 22. 

In cupelling, 369. 

In Pattinson's process, 300, 367. 

In Parkes' process, 367. 
Black slag from the slag-eye furnace, 

analysis of, 124. 
Blake: 

On gaseous fuel in the blast-furnace, 
153. 

On the cyanide assay of copper, 242. 

On the silver bullion assay, 394. 
Blake ore crusher, the 67. 
Blast apparatus : 

For the blast-furnace, 194. 

For the cupelling furnace, 389. 
Blast-furnace : 

Action of fluxes in the, 138. 

Blowers for the, 194. 

Calculation of charge for the, 209. 



INDEX. 



399 



Blast-furnace: 

Capacity of the. 181. 

Chemistry of the. 205. 

Cost of smelting in the, 293. 

Fine ores in the, 229, 239. 

Foundation of the, 179. 

Fuels used in the, 150. 

Influeneeof foreign matter inthe, 138. 

Irregularities in the, 235. 

Lead smelting in the, 132. 

Management of the, 219. 

Matte-smelting in the, 262. 

Operations in the, 219. 

Ores suited to the, 132. 
t Reducing agents in the, 205. 

Slags of the. 133, 135. 

With closed top, 186. 

With open top. 184. 
Blast-pipe, form of the ordinary, 196. 
Blast, pressure of: 

In blo\ving-in. 221. 

In the blast-furnace, 181. 195. 

In the cupelling furnace, 374. 
Blende : 

Action of, in the blast-furnace, 146. 

Influence of. in the roasting and re- 
action process, 83. 

Influence of, in roasting ores, 155. 

Separating of. from lead ore, 147. 
Blow, analysis of lead ore, 37. 
Blowers for the blast-furnace, 194. 
Blow-holes, 236. 
Blowing-in : 

Of the blast furnace, 219. 

Charges, 226. 
Blowing-out of the blast-furnace, 241. 
Blue powder: 

From retorting, 354, 356, 363. 

From the ore-hearth, 128. 
Bone-ash : 
For cupel-test. 390. 
Use of, in fining silver, 392. 
Books, kept in : 

Smelting works, 47, 68, 243. 

Refining works, 366. 
Boshes in blast-furnaces, 182. 
Bottoms from concentrating matte, 

266, 360. 
Bouhv, analysis of galena ore, 87. 
Brasque, 187, 264, 320. 
Breast of the blast-furnace, 191. 
Bricking flue-dust, 291. 
Bricks, number of : 

In an oblong blast-furnace, 178. 

In a roasting-furnace with fuse-box, 
164. 
Brick-work, used in lining cupel-tests, 

391. 
Bridgman's : 

Automatic sampler, 58. 

Laboratory sampler, 64. 



Bridgman's : 

Mixer and divider, 64. 
Brightening of silver, the, 375. 
Brown's portable assay-furnace, 69. 
Browse from the ore-hearth, 112. 
Bruckner cylinder : 

Roasting matte in the, 261. 

Roasting ore in the, 161. 
Brunton's : 

Automatic sampling machine, 55. 

Quartering shovel, 50. 



O. 



Cadmium, effect of, on lead of com- 
merce, 23. 
Cahen on lead-smelting in the rever- 

beratorv furnace, 108. 
Calcining. See Roasting. 
Calcium sulphate, action of, in the 

blast-furnace, 145. 
Calcium sulphide in blast-furnace slags, 

141. 
Calculation of smelting-charges, 74. 
Calculation of the 
Average assay, 64. 
Blast-furnace charge, 209. 
Value of an ore, 75. 
California, lead ores of, 44. 
Canbv on the determination of arsenic, 

* 252. 
Carbonate of lead. See Lead Carbonate. 
Carbonate ores, 25, 27, 28. 
Carbonic acid for refining zinc-bear- 
ing lead, 342. 
Carinthian method of smelting lead ores, 

84. 
Carnellv- Williams on the boiling point 

of lead, 15. 
Carpenter on gold in matte, 254. 
Cast-iron water jackets, 189. 
Cast-steel water jackets, 193. 
Cement as filling material of cupel- 
tests, 390. 
Cerussite, 27. 
Chalcopyrite : 

Action of, in the blast-furnace, 145. 
Effect of, on zinc in the blast-fur- 
nace, 146. 
Influence of, in the roasting and re- 
action process, 82. 
Chamber-dust, 281. 
Changes of galena in the English re- 

verberatory furnace, 100. 
Channelling, 52. 
Charcoal : 

Blowing-in, with, 221. 
In the blast-furnace, 151. 
Charge : 
Calculation of a, for the blast-fur- 
nace, 209. 



400 



INDEX. 



Charge : 

Calculation of a, for the roasting fur- 
nace, 159. 
Descent of, in the blast-furnace, 235. 
Importance of the size of, in the 

blast-furnace, 230. 
Thickness of, in the reverberatory 

roasting-furnace, 159. 
Thickness of, in the reverberatory 
smelting-furnace, 108. 
Charges : 
Made for sampling ores, 67. 
To be made for smelting in the blast- 
furnace, calculation of, 74. 
Charging-floor. work on the, 227. 
Chemistry of the blast-furnace, 205 
Chenhall on separating zinc from lead 

ore, 148. 
Chimney in roast heaps, 255. 
Chisholm, manganiferous iron ores, 140. 
Church on manganese in lead slags, 

140. 
Classification of smelting methods, 78. 
Claudet, analysis of lead ore, 41. 
Clausbruch on f using-charges for roast- 
ing furnaces, 158. 
Clausthal, drawing off of blast-furnace 

fumes, 174. 
Clay as filling material of cupel-tests, 

372, 390. 
Coke : 

As blast-furnace fuel, 150. 
Blowing-in with, 225. 
With anthracite, bituminous coal, 
and charcoal in the blast-furnace, 
153. 
Cobalt: 
Effect of. on lead of commerce, 23. 
Present in matte, 269. 
Colorado : 

Lead ores of, 33. 
Roasting furnaces of, 162. 
Colorado Smelting Company, stall- 
roasted matte from the, 260. 
Color of lead, 14. 
Color of lead oxide (litharge), 15. 
Color of lead silicates, 17. 
Color of lead slags, 138. 
Comparison of cupellation methods, 

394. 
Comparison of sampling methods, 

62. 
Comparison of smelting in the ore 
hearth and the reverberatory fur- 
nace, 110. 
Composition of lead slags, 133. 
Concrete for dust flues, 291. 
Condensation of flue dust, 282. 
Condensers of retorts, 353. 
Conductivity of lead for electricity and 
heat. 15. 



Conductivity of lead oxide for elec- 
tricity, 16. 
Consolidated Kansas City S. & R. Co. : 

Refined lead, 21. 

Hunt & Douglas process at the works 
of the, 268. 
Continuous mechanical sampling, 52. 
Cooling of: 

Flue dust, 282. 

Lead in a kettle, 335. 
Copperas, use of, in bricking flue dust, 

291. 
Copper : 

Effect of, on lead in the blast-fur- 
nace, 146. 

Effect of, on lead of commerce, 22. 

Effect of, on the assay of base bul- 
lion, 249. 

In cupelling. 369. 

In zinc-desilverization, 312. 

Methods of assay of, 251, 271. 
Copper ores, action of, in the blast-fur- 
nace, 145. 
Cordurie. method of refining lead, 341. 
Cost of: 

An oblong blast-furnace, 179. 

A roasting furnace with fuse-box, 164. 

Desilverizing by Parkes' process. 305. 

Desilverizing, the relative, 297. 

Roasting in a furnace with fuse-box, 
165. 

Smelting a neutral ore, 74. 

Smelting in the blast-furnace, 294. 
Crocoite, 29. 
Crooke process, the, 267. 
Croselmire, separating zinc from lead 

ore, 148. 
Crucible-assay: 

For silver in ores, 70. 

For silver in slags, 72. 
Crucible-castings of the blast-furnace, 

186. 
Crucible-filling with lead in the blast- 
furnace. 224. 
Crucibles for lead assays, 70. 
Crucible- walls of the blast-furnace, 187. 
Crusts formed in blowing-in, 222. 
Crusher, the Blake, 67. 
Crystals of : 

Galenite, 25. 

Lead, 14. 

Lead slags, 137. 

Litharge (lead oxide), 15. 
Cupel bottom, method of working, 364. 
Cupellation, 369. 

Comparison of methods, 394. 

Comparison with other desilverizing 
processes, 296. 

The English, 380, 389, 391. 

The German, 371. 

The products of, 376, 401. 



INDEX. 



401 



Curtis, refining of silver, 378. 
Cutting-out of wall accretions. 236. 
Cyanide assay of copper, the, 271. 



Darby, drawing off of blast-furnace 

fumes, 185. 
Davies on desilverizing speise, 250. 
Decomposition of: 

Lead silicates, 17. 

Slags, 274. 
Decopperization of lead by zinc, 311. 
Denver Fire Clay Co., crucibles of the, 

70. 
Depth of the blast-furnace crucible, 

187. 
Descent of the charge in the blast- 
furnace, 235. 
Desilverization of base bullion, 296. 

By cupellation, 369. 

Bv Luce-Rozan's process. 302. 

Bv Parkes' process, 310, 325, 333, 
346. 

By Pattinson's process, 298. 
Desilverization of: 

Matte, 267. 

Speise, 250. 
Desilverizing kettles, 329. 
Desulphurization of galena, 155. 
Determination of: 

Alumina, 277. 

Baryta. 275, 277. 

Iron, 276. 

Lime, 277. 

Magnesia, 277. 

Manganese, 276. 

Silica, 275. 

Sulphur, 273. 

Zinc, 277. 
Determination. See Assaying. 
Devereux, tuyere-box of. 200. 
Dezincification of lead, 339. 
Diaphaneity of lead-slags, 138. 
Dilatation of lead by heat, 15. 
Discharge of lead-kettles, 331. 
Distillation of lead, 15. 
Distillation of zinc-crusts, 348, 354. 
Distribution of impurities, silver and 

gold in base bullion, 244. 
Distribution of the charge on the feed- 
floor, 228. 
Divider and mixer of Bridgman, 64. 
Dobers, analysis of flue-dust, 105. 
Dolomite, effect of, in the blast-fur- 
nace, 141. 
Dolomite, influence of, in the roasting 

and reaction process, 82. 
Double-hearth roasting-furnaces, 160. 
Douglas, calcining-furnace of, 269. 
26 



Drawing off of blast-furnace fumes, 172,. 

174, 184. 
Dressing lead-ores, 25, 28. 
Dross from base bullion, 315, 323, 357, 

376. 
Dry concentration, literature on, 28. 
Dry condensation of flue-dust, 282. 
Drving of the blast-furnace crucible, 

219. 
Drving of the cupel-test. 391. 
Ductility of lead, 14. 
Dump, importance of the, 169. 
Dump, work on the, 234. 
Dumping-car of Nesmith, 202. 
Dust-chambers of blast-furnaces, 282. 
Dust-chambers of roasting furnaces, 164. 
Dwight on coke and anthracite in the 

blast-furnace, 153. 
Dziegiecki, analysis of flue-dust, 105. 



E. 



Edelmann on enriching zinc-crusts, 3*28. 

333. 
Edelmann on zinc retained bv lead, 

24. 
Eilers on lead-slags, 134, 135. 
Eilers, tuyere-box of, 198. 
Elbers on alumina in blast-furnace 

slags, 144. 
Electricity: 

In condensation of flue-dust, 282, 

288. 
In desilverization, 296. 
In treating zinc-crusts, 329. 
Electrolytic assay of copper, 272. 
Endemann on the effect of bismuth 

on lead of commerce, 22. 
Engis, analyses of products from, 89. 
Engis, lead-smelting in the reverber- 

atory furnace at, 87. 
English cupellation, 380, 389, 391. 
English method of lead-smelting in the 

reverberatorv furnace, 90, 100. 
Ems, dust-flue of, 283. 
Ems, results of desilverization at, 297. 
Eureka, Nev., desilverization at, 308. 
Eureka, Nev., working speise at, 250. 
Excavation of blast-furnace founda- 
tion, 179. 
Extraction of silver from lead, 296. 



F. 



Faber du Faur furnace, 349. 
Fan-blowers, 194. 

Feed-holes in the blast-furnace, 184. 
Feeding-down of the blast-furnace, 
236. 



402 



INDEX. 



Feeding the blast-furnace, the manner 

of, 229. 
Feeding of the blast-furnace, the me- 
chanical. 186. 
Ferrous sulphate, its use in bricking 

flue-dust. 291. 
Filling the blast-furnace. 220. 
Filtering of flue-dust, 282. 287. 
Fine ore: 

Agglomerating of, 157, 293. 
Bricking of. 291. 

Effect of, in the blast-furnace, 239. 
Finishing the sample. 62. 
Fire-tops in the blast-furnace, 236. 
Fixed test-supports. 386. 
Flach. process of, for treating zinc- 
crusts. 348. 
Flechner on working hearth-accretions, 

280. 
Flintshire furnace, the, 91. 
Flow of litharge in cupelling, 375. 
Flue-dust, collecting of. 281. 
From cupelling, 391. 
From roasting. 167. 
From smelting in the blast-furnace. 

81, 281. 
From smelting in the ore-hearth. 121. 
From smelting in the reverberatory 

furnace, 105. 
From softening base bullion. 325. 
Manner of treating, 291. ■ 
Fluegger, analysis of lead ore, 36. 
Fluorspar: 

Action of, in the blast-furnace, 112. 
Fcehr on its effects, 142. 
Influence of. in the roasting and re- 
action process, 82. 
Fluxes : 

Action of. in the blast-furnace, 138. 
Purchasing of. 76. 
Receiving and sampling of. 68. 
Fluxing properties of lead oxide 

(litharge). 16. 
Fo?hr on the effect of fluorspar. 142. 
Foreign matter, influence of. in the 

blast-furnace. 138. 
Form of book for : 
Assays, 243. 
Furnace records, 243. 
Receiving ores, 47. 
Record of bins and beds, 68. 
Zinc-desilverization. 366. 
" Fortschaufelungsofen,*' the. 160. 
Foundation of the blast-furnace. 179. 
Fractional selection, the, 151. 
Fraser and Chalmers slag-pots, 201. 
Fraser and Chalmers test-carriage, 386. 
Freeland. analysis of lead ore, 37. 
Freiberg, cooling of flue-dust. 284. 
Freiberg, smelting of roasted matte. 
265. 



Freudenberg. plates for settling flue- 
dust. "286. 
Fuel best suited for the blast-furnace, 

150. 
Fuel, consumption of, in the: 

Blast-furnace, 150, 154. 

Luce-Rozan process, 309. 

Ore-hearth, 119. 

Refining furnace, 341. 

Retorting furnace, 356. 

Roast-heap, 256. 

Roasting reverberatory furnace with 
fuse-box, 165. 

Roasting reverberatory furnace with- 
out fuse-box, 166.* 

Roast -stall, 260. 

Smelting reverberatory furnace, 106. 

Softening furnace, 321* 

Zinc desilverization, 317. 
Fuel consumption under the kettles in 

zinc-desilverizing, 337. 
Fuels : 

Purchasing of , 76. 

Receiving and sampling of, 68. 
Fumes, drawing off of, in the blast-fur- 
nace, 181. 
Fumes. See Chamber-dust, Flue-dust. 
Furnace : 

Accretions, 278, 279. 

Books, form of, 213. 

Cleanings, 281. 

Floor, work on the. 231. 

Gases, analvsis of, 207. 

Products, 244. 

Refuse. 281. 

Site, selection of. 168. 

Sows, 279. 
Furnace, the 

Crucible, for retorting, 348. 

Reverberatory, for cupelling, 371, 380. 

Reverberatory, for liquating, 332. 

Reverberatory, for refining lead, 339. 

Reverberatory, for roasting, 160. 

Reverberatory, for smelting, 79. 

Reverberatorv, for softening base bul- 
lion, 317. 
Fuse-box in roasting-furnaces, 162. 
Fusion of flue-dust, 293. 
Fusibility of : 

Lead, 15. 

Lead silicates, 17. 

Lead-slags, 136. 

Lead sulphide, 18. 

G. 

Galena, 25. 
Assay of, 69. 

Associated minerals of, 26, 
Concentration of, 25. 
Conductivity of, for electricity, 18. 



INDEX. 



403 



Galena: 

Decomposition of, 18, 20. 

Roasting of, 20, 155. 
Gaseous fuel, use of, in the blast-fur- 
nace, 153. 
General arrangement of a : 

Plant for Parkes' process, 313. 

Roasting plant, 164. 

Sampling plant, 67. 

Smelting plant, 169. 
General arrangement of dust-chambers, 

288. 
General arrangement of works : 

Globe S. & R, Company, 170. 

Grant and Omaha S. & R. Company, 
171. 

Montana Smelting Company, 172. 
General reduction process, 132. 
German cupellation, 371.. 
Germania works, Bruckner cylinder at 

the, 161. 
"Gestiibbe." See Brasque. 
Glenn, on feeding the blast-furnace, 229. 
Globe smelting works : 

Bag process at the, 126, 288. 

General arrangement of the, 170. 
Gold : 

Assaying, 72. 

Behavior of, in cupelling, 370. 

Crust, 336, 337, 357. 

Distribution of, in base bullion, 246. 

Occurrence of, in carbonate ores, 29. 

Occurrence of, in galena, 27. 

In the hearth-bottoms, 325. 

In matte, 254. 

Yield of, in Parkes' process, 365. 
Grab sample, the, 69. 
Gray-slag from the roasting and reac- 
tion process, 81. 
Grant and Omaha Smelting Works, 

general arrangement of, 171. 
Gr'uner on lead-smelting in the rever- 

beratory furnace, 108. 
Guyard on : 

Calcium sulphide in blast-furnace 
slags, 141. 

Feeding the blast-furnace, 229. 

Leadville speise, 249 . 

The chemistry of the blast-furnace, 
205. 

The use of lime against wall-accre- 
tions, 238. 



Hagen, water-cooled dust-flues, 283. 
Hahn: 

On alumina in blast-furnace slags, 
143. 

On blowing-in without lead, 225. 

On lead-slags, 133, 135. 



Hahn : 

On the consumption of fuel in the 

blast-furnace, 154. 
On the number of tuyeres in the 
blast-furnace, 195. 
Hammers. See Tools. 
Hampe on the effect of antimony, bis- 
muth, and copper on lead of 
commerce, 22, 23. 
Harbordt on magnesia and manganese 

in lead-slags, 140, 141. 
Hard lead: 
Analyses of, 361. 
Liquating and poling of, 360. 
Softening of, 314. 
Hardness of lead, 14. 
Hartz Mountains: 
Analyses of lead, 21. 
Separating zinc from lead ore, 148. 
Smelting of roasted matte, 262. 
Headden : 

On the effect of altitude on the con- 
sumption of fuel in the blast- 
furnace, 154. 
On the zinc-assay, 355. 
Heap-roasting of 
Matte, 255. 
Ore, 156. 
Hearth-accretions, 240, 279. 
Hearth-bottom from the roasting and 

reaction process, 81. 
Hearth of the blast-furnace, 186, 187. 
Hearth of the reverberatorv furnace for 
Cupelling, 372, 390. 
Roasting, 162, 166, 261. 
Softening, 317. 
Smelting litharge, 362. 
Smelting ore. 84, 87, 91, 100. 
Height of the blast-furnace, 180. 
Height of the blast-furnace above sea- 
level, its effect on the consump- 
tion of fuel, 154. 
Height of the stack of blast-furnaces, 

291. 
Height of the stack of stalls, 257. 
Henri ch : 

On alumina in blast-furnace slags. 

143. 
On blowing-in. 225. 
On feeding the blast-furnace. 230. 
On lead-slags, 133. 

On the action of pyrite in the blast- 
furnace, 145. 
Hering on flue-dust, 287. 
Hermann, comparison of desilverizing 

methods, 296. 
History of lead, 11. 
Hodges on offsets in the roasting- 

hearth, 162. 
Honsell on the melting-point of lith- 
arge, 16. 



404 



INDEX. 



Horizontal section of the blast-furnace, 

180. 
Howe : 
On alumina in blast-furnace slags, 

143. 
On pressure-blowers and fans. 194. 
Hunt and Douglas process No. 2, 268. 



Idaho, lead ores of, 43. 
lies: 

On lead-slags, 132, 135, 137, 138. 

On slag-analysis, 276. 

On the distribution of the charge on 
the feed-floor, 228. 

On wall accretions, 236. 
Iles-Keiper, slag pot of, 20 1 . 
Illing on the amount of zinc required 

in desilverizing, 325. 
Illinois, lead ores of, 31. 
Impurities in base bullion : 

Distribution of, 244. 

Their effect on the assay, 249. 
Impurities, their effect on the value of 

ore, 73. 
Indications of good working: 

On the feed-floor, 230. 

On the furnace-floor, 234. 
Influence of foreign matter in 

The blast-furnace. 138. 

The ore-hearth, 112. 

The reverberatory-furnace, 81. 
Intermediary crystals. 299. 
Intermittent mechanical sampling, 54. 
Iowa, lead ores of, 31. 
Iron: 

Action of, in the blast-furnace, 138. 

And lead sulphide, 18. 

Available for fluxing, 211. 

Determination of, 276. 

Effect of, on lead of commerce, 23. 

Fluxes, analyses of, 139. 

In lead-slags, 136. 
Iron ores, manganiferous, analyses of, 

140. 
Iron ring, use of, in ladling lead, 360. 
Iron vitriol, its use in bricking flue- 
dust, 291. 
Iron-work : 

Of an oblong blast-furnace, 178. 
Of a roasting-furnace with fuse-box, 
164. 
Irregularities in the blast-furnace, 235. 



Jaw crusher, the Blake, 67. 
Jossa on barium sulphide in blast-fur- 
nace slags, 144. 



K. 

Karmarsch on the tenacity of lead, 14. 
Karsten on desilverizing by means of 

zinc, 310. 
Kedzie, analysis of lead ore, 33. 
Keith on electricity in desilverizing 

base bullion, 296. 
Kellar, analyses of lead ore, raw and 

concentrated, 38. 
Keller on desilverizing lead-slags, 273. 
Kempf, Nenninger & Co. : 

On the distribution of gold in base 

bullion, 246. 
On sampling base bullion, 247. 
Kerl: 

Hearth-material for cupelling fur- 
naces, 379. 
Scorification charges in assaying for 
silver, 72. 
Kettle-book, the, 366. 
Kettle-dross, working of, 362. 
Kettles for : 
Desilverizing, 329. 
Liquating zinc-crusts, 331, 338. 
Refining lead, 341. 
Softening base bullion, 317. 
Kettles, the breaking of, 338. 

The life of, 330, 342. 
Kiliani : 

On the conductivity for electricity 

of galena, 18. 
On the conductivity for electricity of 
lead sulphate, 19. 
Kiln-roasting of matte, 260. 
Kiln -roasting of ore, 156. 
Kirchoff on zincing impure base bul- 
lion, 311. 
Krom on dry concentration, 28. 
Kurnakoff on barium sulphide in blast- 
furnace slags, 144. 



Labor required in: 

Cupelling, 377, 401. 

Desilverizing with Luce-Rozan's pro- 
cess, 308. 

Desilverizing with Parkes' process, 
337, 347. 

Retorting, 356. 

Roasting in the reverberatorv fur- 
nace, 165, 166. 

With the blast-furnace on the dump, 
235. 

With the blast-furnace on the feed- 
floor, 230. 

With the blast-furnace on the fur- 
nace-floor, 234. 

With the ore-hearth, 119. 

With the reverberatory-furnace in 
smelting ore, 106. 



IXDEX. 



405 



Laboratory sampler, the Bridgman, 

04. " 
Ladles. See Tools. 

Ladling hard lead. 300. 
Landsberg on the effect of antimony, 
iron, silver, and zinc on lead of 
commerce, 22, 23, 24. 
Latent heat of lead, 15. 
Lautenthal. steaming zincv lead at. 

341. 
Leaching carbonate ores, 28. 
Lead analysis. See Analysis. 
Lead : 

Amount of, from retorting zinc- 
crusts, 350. 
Assaying of, 71. 270. 
Blowing in without, 225. 
Carbonate, 21. 
Changing of, in the blast-furnace 

crucible, 227. 
Decopperization of. with zinc, 311. 
Dezincification of. 339. 
Distribution of, in roasted matte. 

254. 
Extraction of, from srrav ?lags, 87, 

100, 104. 
Extraction of silver from. 290. 
Finely divided, pressed into a solid 

or liquid mass, 14. 
Fume. See Flue-dust. 
History of, 11. 

Level of, in the blast-furnace, 232. 
Liquated from zinc-crusts, 338. 
Loss of. See Yield of. 
Markets and prices of, 13. 
Melting down in the blast-furnace 

crucible. 224. 
Mould on wheels. 344. 
Of commerce, its impurities and their 

effects, 211. 
Ores : 

Distribution of, 25. 
Metallurgical treatment of, 77. 
Of the Atlantic coast. 30. 
Of the Mississippi vallev, 31. 
Of the Pacific coast, 40." 
Of the Rocky Mountains, 33. 
Oxide, properties of, 15. 
Percentage of, in blast-furnace 

charges, 209. 
Poling of, 340. 
Properties of. 14. 
Pump, the Posing, 347. 
Purification bv crvstallization, 298. 
Refining of. 339. 
Silicates, properties of, 10. 
Lead silver ores : 
Assaying of, 09. 
Purchasing of, 73. 
Dead siphon, the Steit-z, 331, 339, 343. 
353. 



Lead slags: 
Composition of, 133, 136. 
Desilverizing of, 273. 
Physical properties of, 136. 
Rich in manganese, 140. 
Types of, 135. 

Used in smelting matte, 265. 
Lead smelting 
In the blast-furnace, 132. 
In the ore- hearth, 110. 
In the ore-hearth, table of results, 

119. 
In the reverberatory-furnace, 79. 
In the reverberatory-furnace, com- 
ments by Cahen and Gruner. 108. 
In the reverberatory furnace, table of 
results, 100. 
Lead speise. See Speis?. 
Lead, statistics of, 11, 
Lead, steaming in kettles, 334. 340. 
Lead subsulphides, 17. 
Lead sulph-antimonites, arsenites, bis- 
muth ites, 29. 
Lead sulphate, properties of, 19. 
Lead sulphide : 

And other metallic sulphides, 18. 
Properties of, 17. 
Roasting of, 20, 155. 
Lead-tap of the blast-furnace, 186. 
Lead-well, the, 187. 

Clogging up of the, 241. 
Lead, yield of, in desilverizing, 297. 
Lead, yield of, in the : 
Blast-furnace. 293. 
Luce-Rozan process. 309. 
Ore-hearth. 119. 
Parkes process, 305. 
Reverberatory furnace, 106. 
Leak, effects of a, in a water jacket, 241. 
Le Chatelier on the 
Chilling of slags, 274. 
Melting-point of lead, 15. 
Length of roasting furnaces, 100. 
Letrange, separating zinc from lead 

ores, 148. 
Lewis and Bartlett. the bag-process of, 

124. 
Lime : 

Action of. in the blast-furnace, 140. 
Determination of, 277. 
Effect of, on slag-crystals. 137. 
Effect of. on the consumption of fuel 

in the blast-furnace, 154. 
In lead-slags. 130. m 

Magnesia and zinc oxide in lead-slags, 
141. 
Lime, use of : 

In bricking flue-dust, 291. 
In removing wall accretions, 238. 
In the roasting and reaction process, 
80. 



400 



INDEX. 



Lime, use of: 

In the refining furnace, 340. 

In the softening furnace, 323. 
Limestone : 

Analyses of, 142. 

Influence of, in the roasting and re- 
action process, 82. 
Life of kettles in desilverizing base bul- 
lion, 330. 

In refining lead, 342. 
Life of retorts, 356. 
Lining, the, of the blast-furnace, 182. 
Liquated lead, adding of, to the kettle, 

336. 
Liquating of zinc-crusts: 

Apparatus, 331. 

Lead obtained from, 338. 
Liquidity of lead-slag, 136. 
Litharge: 

Flow of, in cupelling, 375. 

Properties of, 15. 

Working of, 363. 
Livingstone, lead-slag of, 135. 
Losses in : 

Cupelling, 377, 394. 

Desilverizing, 297. 

Roasting, 158. 

Smelting in the blast-furnace, 293. 

Smelting in the ore-hearth, 119. 

Smelting in the reverberatory fur- 
nace, 106. 
Losses in the 

Luce-Rozan process, 309. 

Parkes process, 365. 
Lowe, separating zinc from lead ore, 

148, 
Luce and Rozan process, the, 302. 
Lumaghi, separating zinc from lead 

ore, 147. 
Lustre of lead, 14. 

Lead-slags, 138. 
Lynch, cupel-test of, 387. 

M. 

Machinery, the placing of, in works, 170, 

171, 172. 
Macintosh on the composition of matte, 

254. 
Magnesia : 

Action of, in the blast-furnace, 140, 

141. 
Determination of, 277. 
Manner of figuring, 141, 211. 
'Magnetism of lead-slags, 138. 
Malleability of lead, 14. 
Management of : 

Cupelling furnaces, 372, 389. 
Ore-hearths, 116. 

Reverberatory furnaces in roasting, 
164, 166, 262. 



Reverberatorv furnaces in smelting, 
86. 88, 89, 97, 99, 102. 

Roast-heaps, 255. 

Roast-stalls, 259. 
Management of the 

Blast-furnace, 219. 

Luce-Rozan plant, 302. 

Parkes plant, 314, 322, 333, 338. 339- 
346, 354, 357. 

Pattinson plant, 300. 
Manganese : 

Action of, in the blast-furnace, 139. 

Determination of, 276. 

Effect of, on lead of commerce. 24. 

Manner of figuring, 139. 
Manufacture of slag-brick, 258. 
Markets of lead, 13. 
Massicot, 15. 
Matte : 

Analyses of, 253. 

Conductivity of, for electricity. 18. 

Desilverization of, 267. 

Heap-roasting of. 255. 

Kiln-roasting of, 260. 

Nickel-cobalt in, 268. 

Reverberatory-roasting of, 261. 

Roasted, lead and silver in. 254. 

Roasting of, 254. 
Matte-slags, 265. 

Matte, smelting of, in the blast-fur- 
nace, 262. 

In the reverberatory furnace. 2(55. 
Matte, stall-roasting of, 256. 
Maxwell-Lyte, separating of zinc from 

lead ore, 148. 
Mechanical sampling: 

Continuous, 52. 

Intermittent, 54. 
Mechernich, double-hearth roasting- 

furnace at, 160. 
Melting-down of 

Base-bullion samples, 247. 

Zinc on base bullion, 333. 
Melting-point of lead oxide (litharge), 

16. 
Melting-points of silver-leads, 298. 
Merchant kettle, moulding lead from 

the, 345. 
Metallurgical treatment of lead ores, 

77. 
Methods of smelting, classification of,. 

78. 
Meyer on the cooling of lead, 335. 
Mimetite, 29. 
Mine La Motte : 

Nickel-matte at, 269. 

Roasting-furnaces at, 165. 
Mississippi Valley : 

Lead ores of, 31 . 

Percentage of ore worked in various 
furnaces, 89. 



INDEX. 



407 



Missouri, lead ores of, 32. 
Mixtures for : 

Lead-assays, 71. 

Silver-assays, 72. 
Moti'et ore-hearth, the, 114. 
Moisture: 

Determination of, 46. 

Sample of ores, 45. 
Monell on roasting-furnaces, 165. 
Monier system, the, for dust-flues, 291. 
Montana, lead ores of, 40. 
Montana Smelting Co. : 

Blast-furnace of the, 177. 

Dust-chambers of the, 289. 

General arrangement of works of the, 
172. 

Tuyere-box of the, 198. 
Moulding lead, 343. 
Moulds for lead on wheels, 344. 
Miiller on the supporting of kettles, 

330. 
Miinster on the composition of matte, 

254. 
Murray : 

Lead-slag of, 135. 

Method of calculating a charge, 212. 

Slag-pot of, 202. 

Tuyere-box of, 199. 
Mrazek on the effect of antimony on 
alloys of nickel (cobalt) and lead, 



N. 



Napier on the effect of bismuth on lead 

of commerce, 22. 
Neill: 
On coke and bituminous coal in the 

blast-furnace, 152. 
On nickel-matte, 269. 
Nesmith, dumping-car of, 202. 
Neutral ore, definition and cost of 

smelting, 74. 
Nevada, lead ores of, 40. 
Newberry, analysis of lead ore, 42. 
New England, lead ores of, 30. 
Newhouse : 

On the calculation of blast-furnace 

charges, 212. 
On the loss of metal in roasting, 157. 
New Mexico, lead ores of, 40. 
New York, lead ores of, 30. 
Nickel : 
Behavior of, in the Pattinson pro- 
cess, 300. 
Effect of, on lead of commerce, 23. 
In speise, 249, 250. 
Matte, 269. 
Nitre, use of, in fining silver, 393. 
Nolte on the reduction of lead sulphide 
by iron, 18. 



Non-argentiferous lead ores, purchas- 
ing of, 75. 

Nose formed in the blast-furnace, 231, 
234. 



Oil, use of, as fuel in reverberatorv 

furnaces, 322, 353. 
Old retorts, working of, 364. 
Omaha and Grant Works : 

Dust-flues of, 291. 

General arrangement of, 171. 
Open stalls, 256. 
Ore-fines trickling through the charge, 

239. 
Ore-hearth : 

Blue powder from the, 128. 

Comparison with the reverberatory- 
furnace, 110. 

Influence of foreign matter in the, 
112. 

Lead smelting in the, 110. 

Management of the, 116. 

Pattinson on the, 117. 

Products of the. 111. 

Recovery of flue-dust from the, 124. 

Slags from the. 118. 

Table of results, 1.19. 

The American water-back, 113. 

The Moffet, 114. 

The Rossie. 114. 

The Scotch, 112. 

Treatment of slags, 120. 
Ore-sampling machines, 52. 
Ores, assay sample, the, 47. 

Carbonate and sulphide, 45. 

Form of receiving-book. 45. 

Form of record of bins and beds, 68. 

Moisture sample of, 45. 

Of lead. See Lead Ores. 

Receiving and weighing of, 45. 

Size of, in roasting and smelting in 
the reverberatorv-furnace, 79. 
159. 

Suited for the blast-furnace, 132. 

Suited for the roasting and reaction 
process, 79. 

The roasting of, 155. 

The smelting of, 78. 
Outline of the 

Blast-furnace process, 132. 

Crooke process, 267. 

Cupellation process, 369. 
Outline of the Hunt and Douglas pro- 
cess, No. 2, 268. 

Of the Pattinson process, 298. 

Of the Pattinson-Parkes process, 367. 

Of the Parkes process, 313. 

Of the roasting and reaction process, 
79. 



408 



INDEX. 



Oxide lead-ores, influence of, in the 

roasting and reaction process, 83. 
Oxide of lead. See Lead-oxide. 
Oxides from steaming zinc-bearing 

lead, 341. 
Oxides of iron, influence of, in the 

roasting and reaction process, 82. 
Oxidizing action of lead-oxide (litharge), 

16. 
Oxidizing action of manganese-oxides, 

139. 
Over-fire in the blast-furnace, 236. 



P. 



Pacific, lead-ore of the, 40. 

Page, lead-slag of, 135. 

Pans, iron, for refining and softening 

furnaces, 318. 
Pan-slimes for bricking ore, 292. 
Parnell, separating zinc from lead ore, 

148. 
Pattinson on the ore-hearth, 117. 
Pattinson-Parkes process, the, 367. 
Pattinson process, comparison with 

other processes, 296. 
Pattinson process, the, 298. 
Parkes-Pattinson process, the, 367. 
Parkes' process: 
Comparison with other processes, 296. 
Cost of the, 365. 
Management of the. 314, 322. 333, 

338, 339-346, 354, 357. 
Result from the, 356. 
Tree of the, 364. 
Yield of metal from the, 365. 
Payment for lead, manner of, 73. 
Payment for copper, gold, and silver, 

manner of, 74. 
Pennsylvania Lead Co., analysis of lead 

from the, 21. 
Pearce on manganese in slags, 140. 
Pearce on the determination of arsenic, 

252. 
Percy on wrought iron crucibles in 

lead-assays, 70. 
Peters on alumina in the blast-furnace, 

143. 
Peters on bricking fine ores, 292. 
Pfort, method for drawing off blast- 
furnace fumes, 184. 
Phillips, analysis of lead-ore, 84. 
Physical properties of lead-slags, 136. 
Pietsch, analysis of residue from rever- 

beratory smelting furnace, 105. 
Pipe, form of the ordinary blast, 196. 
Pipe, the ore-sampler, 52* 
Plant of the 

Globe Smelting Works, 170. 
Luce-Rozan Process, 302. 
Montana Smelting Works, 172. 



Plant of the 

Omaha and Grant Smelting Works, 
171. 

Parkes Process, 313. 

Pattinson Process, 300. 
Plattner: 

Cupelling furnace of, 371. 

On lead and silver in roasted matte, 
254. 

On roasting lead-sulphide, 20. 

On the effect of bismuth on lead of 
commerce, 22. 

On tin-skimmings, 361. 

On zinc for desilverizing, 326. 
Poling of lead, 346. 
Polings, 362. 
Precipitation process, the 

In the blast-furnace, 132. 

In the reverberatory furnace, 79. 
Pressure in bricking fines, 292. 

Of blast in blowing-iti, 221. 

Of blast in the blast-furnace, 181, 
195. 

Of blast in the cupelling-f urnace, 374. 
Pribram : 

Analysis of lead from, 21. 

Cupelling furnace of, 371. 

Stall-roasting at, 256. 

The Luce-Rozan process at, 302. 
Price of lead, the, 13. 
Process \ 

The blast-furnace, 132. 

The Crooke, 267. 

The cupellation, 369. 

The general reduction, 132. 

The Hunt and Douglas, No. 2, 268. 

The Pattinson, 298. 

The Pattinson-Parkes, 367. 

The Parkes, 313. 

The precipitation in the blast-fur- 
nace, 132. 

The precipitation in the reverbera- 
tor} 7 furnace, 79. 

The roasting and reaction, 79. 
Production of lead in the world, 12. 
Properties of lead, 14. 
Pueblo, the Crooke process at, 267. 
Purchasing of fluxes and fuels, 76. 

Of lead-silver ores, 73. 

Of non-argentiferous ores, 75. 
Pyrite: 

Action of, in the blast-furnace, 145. 

Influence of, in the roasting and re- 
action process, 82. 

Manner of figuring, 145. 
Pypomorphite, 29. 



Quartering, 48. 

Quartering-shovel, the Brunton, 50. 



IXDEX. 



409 



Quartz, a flux in concentrating matte, 
"266. 
A flux in roasting galena ores, 156, 
166. 



Rabbling in roasting in the reverbera- 
tor)' furnace, 160. 
Raht. lead-slag of, 135. 
Raibl, lead smelting at, 84. 
Ramnielsberg: 
On the composition of matte, 254. 
On the roasting of lead sulphide, 20. 
Rapid smelting with zincy ores, 147. 
Rapp on coke and anthracite in the 

blast-furnace, 153. 
Reactions, special, with lead sulphide, 

20. 
Reducing agents in the blast-furnace, 

205. 
Receiving: 

Base bullion. 314. 
Fluxes and fuels. 68. 
Ores, 45. 
"Record of bins and beds, 68. 
Reduction of: 

Lead sulphate, 19. 

Lead sulphide by copper, iron, and 

zinc, 18. 
Lead silicates, 17. 
Reed, table of largest sizes in sampling 

ore. 47. 
Refined lead: 

Analyses of, 21, 87, 90, 118, 367. 
Moulding of, 343. 
Yield of, in desilverizing, 297. 
Yield of, in the Luce-Rozan process, 

309. 
Yield of, in the Parkes process, 347. 
Refiner on melting base-bullion sam- 
ples, 248. 
Refining: 

Crude silver, 377. 

Desilverized lead, 339. 

Furnace for lead, 339. 

Furnace, moulding lead from the, 

343. 
Kettles. 341. 
Oil used in. 322. 
Oxides from, 341. 
Refining skimmings: 
Amount of, 341." 
Working of, 362. 
Reich : 

On the melting-points of silver-leads, 

298. 
On the percentage of iron in lead, 24. 
On the specific gravity of pure lead, 
14. 



Residue from lead smelting in the re- 
verberatorv furnace, 89, 90, 105. 
Results from cupelling, 377, 394. 
Results from roasting: 

In reverberatorv furnaces, 165, 166, 
262. 

In heaps. 255, 

In kilns. 261. 

In stalls, 259. 
Results from smelting: 

In the blast-furnace, 181, 293. 

In the slag-eye furnace, 122, 128. 
Results from : 

The Luce-Rozan process, 308, 309. 

The Parkes process, 356, 365. 
Results, table of, from lead-smelting: 

In the ore-hearth, 119. 

In the reverberatorv furnace, 106. 
Reverberatory furnace: 

For liquating zinc-crusts, 332. 

For refining zincy lead, 339. 

For roasting matte, 261. 

For roasting ore, 160. 

For smelting ores, 106. 

For smelting refining skimmings, 362. 

Use of oil in the, 322. 
Retorting zinc-crusts, 348, 354, 356. 
Retort-dross, 356, 363. 
Retorts : 

For distilling zinc-crusts, 350, 356. 

Working up of old, 364. 
Revolving cylinder for roasting, 161. 
Rhodes on silver in copperv bullion, 

246. 
Richmond, Xev. : 

Dust-flue of, 283. 

The Luce-Rozan process at, 308. 
Richter on the cooling of flue-dust, 

284. 
Rittinger, a mechanical sampler, 54. 
Roasted blue powder from the ore- 
hearth, 118. 
Roasted matte, smelting of, 262. 
Roasted ore, analyses of, 167. 
Roast-heaps, 255. 
Roast-kilns, 206. 
Roast-stalls, 257. 
Roasting and reaction process, 79. 

Influence of foreign matter in the, 
81. 

Products of the. 81. 
Roasting and reduction process, the, 

132. 
Roasting furnace at Mine La Motte, 

165. 
Roasting furnaces, 160, 261. 

Analyses of products, 167. 

General arrangement of, 164. 

With fuse-box, 162. 

Without fuse-box, 165. 



410 



INDEX. 



Roasting matte, 254. 

In heaps, 255. 

In kilns, 260. 

In reverberatory furnaces, 261. 

In stalls, 267. 
Roasting ores, 155. 

Products of, 166. 
Roberts- Austin on solidifying and li- 
quefying powdered metals by 
pressure, 14. 
Rocky Mountains, lead ores of the, 33. 
Rolker, analyses and assays of lead ores, 

37. 
Rosing : 

Improvements of, in retorting, 356. 

Lead-pump of, 347. 

On Bessemerizing base bullion, 296. 

On the electrolysis of zinc-crusts, 329. 

On the plant for Parkes' process, 313. 

On the theory of Parkes' process, 
310. 

Test-support of, 389. 

Wires for settling flue-dust of, 287. 
Rossie, the ore-hearth at, 114. 
Rossler : 

On enriching zinc-crusts, 333. 

On refining silver, 378. 

On steaming of zincy lead, 342. 

On the assay of base bullion, 249. 

On the amount of zinc retained by 
lead, 24. 
Roswag : 

On the amount of zinc in desilveriz- 
ing, 325. 

On the distribution of silver in base 
bullion, 245. 



S. 



Salamanders. See Hearth Accretions. 
Sample : 

Finishing the, 62. 

Of desilverized lead in refining, 340. 

Of base bullion after drossing, 316. 

Of base bullion after softening, 317. 
Sampler: 

The Bridgman, 58. 

The Brunton, 55. 

The Rittinger, 54. 
Sampling: 

Charges made for, 67. 

Department, general arrangement of, 
67. 

Fluxes and fuels, 68. 

Methods, comparison of, 62. 

Mechanical continuous, 52. 

Mechanical intermittent, 54. 

Mill with Bridgman's sampler, 60. 

Mill with Brunton's sampler, 56. 

Of base bullion, 246. 



Sampling : 

Of fine silver, 393. 

Of retort bullion, 355. 

Of slag, 274. 

Ore, table of largest sizes, 47. 
Sand in fining silver, 378. 
Scales on the charging-floor, 227. 
Schertel : 

Analyses of matte and slag, 265, 266, 

On furnace gases, 207. 

On melting temperature of lead- 
slags, 205. 

On softening dross, 316. 
Schlosser and Ernst, cooling of flue- 
dust, 285. 
Schmieder,on the support of kettles, 342. 
Schneider: 

On alumina in lead-slags, 143. 

On lead-slags, 135. 

On lime in lead-slags, 136, 141. 
Schweder : 

On the composition of matte, 254. 

On the decomposition of barite, 145. 
Scorification-assay : 

For silver in ores, 70. 

For silver in slags, 72. 
Scotch ore-hearth, the, 112. 
Season of the year, effect of, on the 
consumption of fuel in the blast- 
furnace, 154. 
Selection : 

Of a furnace site, 168. 

Fractional, 51. 
Settling of flue-dust, 285. 
Shaft-furnace : 

Matte-roasting experiment, 261. 

See Blast Furnace. 
Shaft of the blast-furnace, 182. 
Shovel-system of charging, the, 227. 
Siemens and Halske, separating zinc 

from lead ore, 148. 
Silesian method of lead-smelting, 100. 
Silica and lead sulphate, 19. 
Silica : 

As flux in concentrating matte, 266. 

As flux in slag-roasting galena, 156, 
166. 

Determination of, 275. 

Influence of, in the roasting and re- 
action process, 81. 

In lead-slags, 136. 
Silicates of lead. See Lead Silicates. 
Silver and gold, behavior of, in the 
roasting and reaction process, 83. 
Silver and zinc, 310. 
Silver : 

Assays, 70, 246. 

Addition of, in slag assays, 71. 

Behavior of, in cupellation, 370. 

Crusts, 337. 

Distribution of, in base bullion, 245. 



INDEX. 



411 



Silver : 

Effect of, on lead of commerce, 22. 

Extraction of, from lead, 296. 

In carbonate ores, 28, 29. 

In flue-dust, 282. 

In galena ores, 26, 27. 

In hearth-bottom of softening fur- 
nace, 325. 

In litharge, 370, 376, 394. 

In raw matte, 254. 

In refined lead, 297, 328. 

In roasted matte, 254. 

In softening dross, 315. 

In softening skimmings, 359. 

In zinc-crusts, 337. 

Its effect on the melting point of 
lead, 298. 

Islet, channelling at, 52. 

Loss in concentrating lead-ore, 26, 28. 

Loss in cupelling, 377. 

Loss in desilverizing, 297. 

Loss in roasting, 158. 

Loss in smelting in the blast-fur- 
nace, 293. 

Loss in smelting in the reverberatory 
furnace, 105. 
Silver, loss in the 

Luce-Rozan process, 309. 

Parkes process, 365. 
Silver ores : 

Crucible-assays of, 70. 

Scorification-assays of, 70. 
Silver sulphate in refining silver, 378. 
Simmonet, separating zinc from lead 

ore, 148. 
Single-hearth roasting furnaces, 160. 
Siphon-tap, the Arents, 187. 
Siphon, the Steitz, 331, 339, 343, 353. 
Size of charge, importance of, in the 

blast-furnace, 230. 
Size of ore for roasting in the reverbera- 
tory furnace, 159. 

For smelting in the reverberatory 
furnace, 79. 
Skimmings from the softening furnace : 

Amount of, 324. 

Working of, 358. 
Skimming zinc-crusts, manner of, 335. 
.Slag. See also Lead Slag. 
Slag: 

' Addition of, to the blast-furnace, 142. 

Method of analyzing, 274. 
Slag-brick, manufacture of, 258. 
Slag-dump : 

Importance of the, 169. 

Management of the, 234. 
,Slag-eve furnace, the, 120, 122. 
Slag from refining silver, 378, 380. 
,Slag from the : 

Ore-hearth, 118. 120. 

Slag-eye furnace, 124. 



Slagging-hearth in roasting furnaces, 

162. 
Slag-lead from Raibl, 87. 
Slag-pots, 201. 

Slag-tap of the blast-furnace, 186. 
Slag, tapping of, from the blast-fur- 
nace, 231. 
Slags, Balling on the compositionof,133. 
Slater, separating zinc from lead ore, 

149. 
Smelting of roasted matte, 262. 
Smelting. See Lead Smelting. 
Sodium sulphate, action of, in the blast- 
furnace, 145. 
Softening base bullion, 314, 322. 
Softening dross, working of the, 357. 

Furnaces, 317. 

Skimmings, working of the, 358. 
Solubility of lead, 15. 

Of lead oxide (litharge), 16. 
South Dakota, lead ores of, 39. 
Sows, 279. 
Specific gravity of lead, 14. 

Of lead slags, 138. 
Specific heat of lead, 15. 
Speise, 149, 249. 

Spezia, Italy, Bruckner cylinder at, 162. 
Split-shovel, the, 51. 
Spray jackets, the, 190. 
Stack, height of : 

For blast-furnaces, 291. 

For stalls, 257. 
Stall-roasting of matte, 256. 

Of ore, 156. 
Starting the smelting in the blast-fur- 
nace, 223. 
Statistics of lead, 11. 
Steam and carbonic acid in refining 

lead, 342. 
Steam : 

Effects of, on zincy lead, 334, 340. 

Use of, in condensing flue-dust, 286. 
Steaming to refine lead, 341. 

To remove additional zinc-crusts, 
338. 

To soften lead, 317. 
Steam-Pattinson process, the, 302. 
Steel : 

Kettles of. 330, 342. 

Water jackets of cast, 193. 
Steitz, lead-siphon of, 331, 339, 343, 

353. 
Stetefeldt, furnace of, for roasting 

matte, 261. 
Stetefeldt on Pattinson's process, 299. 
Stiperstones, England, lead smelting at, 

92. 
Stirring in of zinc, 333. 
Stolzite, 29. 

Stone on alumina in blast-furnace 
slags, 144. 



412 



IXDEX. 



Sulphate of lead. See Lend Sulphate. 
Sulphide of lead. See Lead Sulphide. 
Sulphide ores, 25. 
Sulphides, their effect on roasting lead 

sulphide, 20. 
Sulphur : 

Determination of, 273. 

Percentage of, in roasted matte. 256, 

259, 200, 261. 
Percentage of, in roasted ore. 157. 
Sulphuric acid made from lead matte, 

260. 
Support of desilverizing kettles. 330. 
Swinging pipe for moulding lead, 

346. 
Swinging test-support, 386, 388. 
Systems of crystallization, 299. 



Table for Parkes" process, 364. 
Table of : 
Carbonate ores, 28. 
Blowing-in charges, 226. 
Changes of galena in the English re- 
verberator}- furnace, 100. 
Concentration of silver in the Luee- 

Rozan process, 308. 
Crucible-mixtures for lead ores, 71. 
Crucible-mixtures for silver ores, 

72. 
Fusible compounds of lead oxide with 

metallic oxides, 16. 
Largest sizes in sampling ores. 47. 
Raw and dressed galena ores. 26. 
Results of lead-smelting in the ore- 
hearth. 119. 
Results of lead-smelting in the rever- 
beratorv furnace, 106. 
Scorification- mixtures. 72. 
Slags, by Balling, 134. 
The merits of the desilverization 

methods. 297. 
The world's production of lead, 12. 
Typical lead-slags, 135. 
Zinc additions for gold, 328. 
Tamping-irons, 373, 390. 
Tap-hole, difficulty of opening, 232. 
Tapping jacket, 191. 
Tap. the automatic, of Arents, 187. 
Tarnowitz, Silesia, lead-smelting at, 

101. 
Tatham. retorting-furnace of, 351. 
Tavlor and Brunton. analyses of Aspen 

ores, 39. 
Telescope-stack, the, 185. 
Tenacity of: 
Lead,' 14. 
Lead-slags, 138. 
Temperature in cupelling, 374. 



: Terhune: 

On cast-steel water jackets. 193. 
On roasting matte, 261. 
Sectional slag-pot of, 201. 
Test-rings, 383. 
Test-ring supports, 385. 
Tests for fine silver, 393. 
Texas, lead ores of, 33. 
Thickness of charge : 

In roasting in the reverberatorv fur- 
nace, 159. 160. 
In smelting in the reverberatorv fur- 
nace, 108. 
Thompson on zinc-silver alloys. 310. 
Thum, separating zinc from lead ore, 

147. 
Time a blast-furnace can remain 

banked, 241. 
Time required for: 
Blowing-in, 223. 
Cupelling at Pribram. 376. 
Desilverizing in the Luce-Rozan pro- 
cess, 308. 
Desilverizing in the Parkes process. 

337. 
Heap-roasting, 256. 
Quartering, 50. 

Refining lead in the kettle. 342. 
Refining lead in the reverberatorv 

furnace. 322, 341. 
Retorting. 356. 
Roasting matte in the reverberatorv 

furnace, 261. 262. 
Roasting ore in the reverberatorv 

furnace, 160. 165, 166. 
Smelting a charge in the ore-hearth, 

117. 
Smelting a charge in the reverbera- 
torv furnace, 106. 
Softening, 322. 
Stall-roasting. 259. 
Tin, arsenic, antimony, lead alloy. 361. 
Tin. effect of, on lead of commerce. 23. 
Tin skimmings from softening base 

bullion, 323, 361. 
Tools required : 
At the cupelling furnace. 375. 389. 
At the English reverberatorv fur- 
nace, 97. 
At the desilverizing kettles. 334. 335, 

336. 
At the Moffet ore-hearth, 116. 
At the retorting furnace. 356. 
At the roasting furnace with f.^e- 

box, 164. 
At the Scotch ore-hearth. 116. 
At the Silesian reverberatorv furnace, 

102. 
At the softening furnace, 323. 
On the dump, 235. 
On the feed-floor, 230. 



IXDEX. 



413 



Tools required on the furnace-floor, 234. 
Torrev and Eaton on gold in base bul- 
lion, 246. 
Torrey on regulating the temperature 

in a muffle. 248. 
Transportation of slair, 234. 
Tuyere- boxes. 198. 

Tuyere, appearance of, while the blast- 
furnace is running. 231. 
Tuyeres: 

Number of. in the blast-furnace, 195. 
Position of, in cast-iron water jackets, 

191. 
Position of, in the English cupelling 

furnace, 382. 
Position of, in the German cupelling 

furnace. 371. 
Position of. in the Plattner cupelling 

furnace, 372. 
Position of. in the softening furnace, 
324. 
Types of lead-slags, 135. 



U. 



United States, lead ores of the. 30. 
Upper Mississippi Vallev. lead ores of 

the, 31. 
Utah, lead ores of, 41. 



V. 



Value of ore. calculation of the. To. 
Vanadinite, 29. 

Vertical section of the blast-furnace. 

182. 
Virginia, lead ores of. 30. 
Virgin lead of Carinthia, 87. 
Volatilization of: 

Lead sulphide. 18. 

Lead oxide (litharge), 16. 

Metal in the blast-furnace, 281. 

Zinc in the blast-furnace, 147. 

Zinc in refining lead. 340. 

Zinc in treating zinc-crusts, 348. 
Von Schulz and Low : 

On the assay of lead, 270. 

On the assav of zinc, 277. 



W. 

Wait on silver in litharge, 370. 
Warming : 

Of the blast-furnace crucible. 219. 

Of the cupel-test, 391. 
Wall-accretions, 236, 278. 
Washing galena ore, its effect in the 
loss in silver. 26. 



Water 

Acid, its effect in bricking flue-dust. 
291. 

Amount of. used in jackets, 194. 

Back, the, in the blast-furnace, 190. 

Back, the, in the ore-hearth. 113. 

Box, the, in the ore-hearth, 113. 

Cooled tuyeres, when first used, 190. 

Cooling of fire-bridge, 164. 

Cooling of the refining furnace. 339. 

Cooling of the softening furnace. 319. 

Cooling of test-rings, 383. 

For jackets, circulation, and distribu- 
tion, 193. 

Impure, its effect on the jacket. 194. 

Jackets, 189. 

Jackets, effect of a leak in, 241. 

Jackets, management of, 231. 

Jackets of cast iron. 191. 

Jackets of cast steel, 193. 

Jackets of wrought iron. 192. 

Jackets, temperature of, 231. 

Use of. in condensing flue-dust, 282. 

Use of, in cooling flue-dust, 283. 

Use of, in cooling lead, 307, 335. 
Weighing of ores, 45. 
Weight of charge of the blast-furnace. 

importance of, 230. 
Werner : 

Adjustable tuvere pipe of. 196. 

Slag-pot of, 202. 
West, separating zinc from lead ore, 

148. 
Wet assav of lead-ores, 270. 

Of ores. 73. 
Wet condensation of flue-dust. 282. 
Wet methods of treating lead-ores. 78. 
Wetting down of flue-dust, 291. 
White metal from concentrating lead- 
matte, 266. 
White paint, composition of, 130. 
Williams : 

On a wet lead-assay method. 270. 

Products from the air-furnace. 90. 
Wilson, separating zinc from lead-ore. 

148. 
Wimmer. hearth material for the cupel- 
ling-furnace. 379. 
Wisconsin, lead-ores of. 31. 
Wood, amount needed: 

In the ore-hearth, 119. 

In the reverberatorv roasting fur- 
nace. 166. 

In the reverberatorv smelting fur- 
nace, 106. 

In the roast-heap, 256. 

In the roast-stall, 260. 
Wood, use of: 

In blowing in, 224. 

In removing crusts from the reverber- 
atorv roasting furnace, 161. 



414 



INDEX. 



World, production of lead of the, 12. 
Wrought-iron water jackets, 192. 
Wulfenite, 29. 
Wuth, analysis of lead-ore, 42. 



Yield of metal: 

In the blast-furnace, 293. 
In the Luce-Rozan process, 309. 
In the ore-hearth, 119. 
In the Parkes process, 365. 
In the reverberatory smelting fur- 
nace, 106. 
Yield of refined lead in the Parkes pro- 
cess, 347. 



Zinc and lead, 24. 
Zinc: 

Amount added for extracting gold, 
328. 

Amount added for extracting silver, 
335. 

Amount consumed in desilverizing, 
337. 

Amount saved in retorting, 356. 

Blende, action of, in the blast-fur- 
nace, 146. 

Its influence on the roasting and re- 
action process, 83. 

Its influence on the roasting of ores, 
155. 

Separation of, from lead-ore, 147. 

Crust, analyses of, 327. 



Zinc: 

Crust, liquation of, 338. 

Crust, treatment of, 347. 

Crust, weight of, 337. 

Crust, yield of rich lead from, 356 , 

Decopperization of lead by, 311. 

Desilverization of lead by, 310, 325, 
333, 346. 

Determination of, 277. 

Effect on consumption of fuel in the 
blast-furnace, 154. 

Effect on lead of commerce, 24. 

In slags from smelting refinery by- 
products, 149. 

Manner of figuring, 147, 211. 

Melting down of, in desilverizing, 333. 

Oxide, action of, in the blast-furnace, 
146. 

Oxide, effect of, on slag-crystals, 137. 

Oxide, magnesia and lime in blast- 
furnace slags, 141. 

Required purity for desilverization, 
312. 

Separation of, from lead by means of 
steam, 341. 

Separation of, from lead in the refin- 
ing furnace, 339. 

Silver alloys, 310. 

Stirring-in of, 334. 

Sulphate, use of, in bricking fine ore, 
291. 

Volatilization of, in the blast-furnace, 
147. 

Volatilization of, in refining lead, 340. 

Volatilization of, in treating zinc- 
crusts, 348. 



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XI 



CARL GAIL, Pkest. 

H. UNZICKER, Gen. Mgr. and Sect. 



E. BL* MILLER, V. P. and Treas. 
M. UNZICKER, Western Rep, 



CHICAGO IRON WORKS, 

MANUFACTURERS OF 

MINING * MACHINERY. 



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Sectional Machinery for Transportation on Mnle Back, 

WORKS AND GENERAL OFFICE : 

CLTBOUKN ATE. and WILLOW ST., CHICAGO, ILL., U. S. A. 

WESTERN OFFICE: 

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XII ADVERTISEMENTS. 



THE METALLURGY OF STEEL 



BY 



HENRY M. HOWE, A.M., S.B. 



Royal Quarto, Handsomely Bound, Printed on Superfine 
Paper, and Profusely Illustrated. 

THIRD EDITION. 

Price, ----- $10.00. 



This work is the most notable contribution to the literature of iron and 
steel metallurgy ever published. The series of papers on the subject which 
have appeared as supplements to the " Engineering and Mining Journal " 
during the past two years have attracted world-wide attention and have re- 
ceived the heartiest commendation from all quarters. The volume now 
published presents this material in much more convenient shape, with con- 
siderable additional matter, giving the results of the most recent research, 
experiment and practice. Mr. Howe also presents a complete review of all 
important conclusions reached by earlier investigators, and his masterly dis- 
cussion of them renders the work classic. Every statement and citation has 
been carefully weighed and verified and the references to the literature of 
the subject are given minutely, the book thus furnishing in itself a key to the 
whole range of steel metallurgy. It also furnishes the results of much new 
and original investigation, specially undertaken for the present work. 

Every metallurgist, every manufacturer of steel in any form, and all who 
are interested in the iron or steel industries, and all engineers who use iron or 
steel should have this standard work and cannot afford to be without it. 



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ADVERTISEMENTS. XIII 



THE 



COLORADO IRON WORKS, 



DENVER, COLORADO, 



MANUFACTURERS OF 



COPPER AND SILVER-LEAD 

SMELTING FURNACES 



MILLING AND MINING MACHINERY 



AND ALL- 



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ARE OUR SPECIALTIES. 



XIV ADVERTISEMENTS. 



THE PHOSPHATES OF AMERICA 

Where and How They Occur; How They Are Minei; 
anil What They Cost. 

With Practical Treatises on the Manufacture of Sulphuric Acid, Acid 
Phosphate, Phosphoric Acid and Concentrated Superphos- 
phates, and Select Methods of Chemical Analysis. 



FRANCIS WYATT, m. r>. 
BOUND IN CLOTH. PROFUSELY ILLUSTRATED. PRICE ONLY S4.00. 



It is the first work of its kind ever published, and can- 
not fail to be welcomed by every one at all interested 
in the subject of phosphates or phosphate mining. 



TO THE MINER 



This remarkable book is indispensable, describing and fully illustrating the 
geological occurrence; the various m idem methods of mining; the actual mining 
cost and the chemical composition of all the workable phosphate deposits of the 
American continent. 

TO THE FERTILIZER MANUFACTURER 

It is, as the title implies, a standard authority in every 
. phase of his industry. It gives methods, resulting from long 
I ractical experience, by vhich the most satisfactory results 
are obtained from all grades of phosphates, including those 
chemically defective and heretofore unused. It also contains 
valuable working tables whereby perfect products can be 
produced by those unfamiliar with chemistry. 



TO THE CHEMIST 



It must prove a veritable guide and reference book, since he will find in it every 
detail required for reliable, rapid, and strictly scientific analyses of all material used, 
in the industries indicated by the title. The methods are those used in the author's 
own laboratory, and are supplemented by detailed descriptions of manipulation and 
apparatus; factors for facilitating calculations; formulae for preparation of all re- 
agents, and instructions for estimating and stating results of analyses. 

TO THE GENERAL READER 

It affords an opportunity, hitherto lacking, of studying a 
concise treatise on phosphates and fertilizers, by a practical 
and conscientious expert in everyday language. The capitalist, 
banker, merchant, and intelligent farmer may all derive from 
it with facility a liberal education on the important subjects 
treated, as the use of scientific terms is restricted to the tech- 
nical pages alone. 



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AIR COMPRESSORS. 




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Cooling with Copper Tubes. Many Late Improvements. 



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XVI ADVERTISEMENTS. 



THE 



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WITH 

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AND 

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TRANSLATED and EDITED 
BY 

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Bound in Cloth, 113 Pages, - Price, $1.50. 



THE SCIENTIFIC PUBLISHING COMPANY, 

Publishers, 

27 PARK PLACE, NEW YORK. 



ADVERTISEMENTS. XV[[ 



Florida, 

South. Carolina 

and Canadian 

PHOSPHATES. 



Giving a Complete Account of Their Occurrence, 

Methods and Cost of Production, Quantities 

Raised, and Commercial Importance. 



— BY — 

C. C. HOYER MILLAR. 



The author gives his experience and investigations during the last few years 
in the Phosphate Fields of Florida, South Carolina and Canada. 



Bound in Cloth, 223 pages ; Price, $2.50. 



TABLE OB> CONTENTS. 

Chapter I. Introduction. 

Chapter II. Florida Phosphates. 

Chapter III. South Carolina Phosphates. 

Chapter IV. Canadian Phosphates. 

Appendix — Analysis of Various Phosphates. 



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27 Park Place, New York. 



XVIII ADVERTISKMENTS. 



MINING ACCIDENTS 



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With Discussion by Leading Experts. Also, the United 

States, British and Prussian Laws relating to 

the Working of Coal Mines. 

Price, - - - $4.00 in Cloth. 



Contents : 
Mining Accidents. By Sir Frederick A. Abel. With discussion by President 
Bruce, of the British Institute of Civil Engineers; and Prof. Arnold Lupton, C. 
Tylden Wright, Emerson Bainbridge, William Morgans, Sydney F. Walker, Col. 
Paget Mosley, Henry Hall, Col. J. D. Shakespear, Stephen Humble, Sir George 
Elliot, Sir Warington Smyth, A. R. Sawyer, A. Giles, R. Bedlington, Edward 
Combes, George Seymour, Henry Harries, William Cochrane, James Ashworth, J. 

B. Atkinson, W. N. Atkinson, Bennett H. Brough, T. Foster Brown, S. B. Coxon, 

C. Le Neve Foster, W. Galloway, Max Georgi, W. S. Gresley, J. A. Longdon, 
A. R. Sennett, M. H. N. Story Maskelyne, Arthur Sopwith, A. L. Steavenson, A. 
H. Stokes and others. 

List of safety appliances, with description of detachment of mineral from its 
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in accessible form. 

Of the unanimously favorable criticisms of this book, we have only space 
to quote one : 

4 ' It is a work that should be in the hands of every intelligent man connected 
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XIX 



CHEMICAL AND GEOLOGICAL 



ESSAYS 



BY 



THOMAS STERRY HUNT, m. a., ll. id., 

Author of " Mineral Physiology and Physiography," "A New Basia 
for Chemistry," " Systematic Mineralogy," etc. 



FOURTH EDITION, 



REVISED AND ENLARGED. 



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TABLE OF CONTENTS. 





Preface ; 


XII. 


I. 


Theory of Igneous Rocks and Yol 






canoes; 


XIII. 


II. 


Some Points in Chemical Geology ; 




III. 


The Chemistry of Metamorphic 






Rocks; 


XIV. 


IY. 


The Chemistry of the Primeval 
Earth ; 


XY. 


Y. 


The Origin of Mountains; 


XYI. 


YI. 


The Probable Seat of Yocanic 






Action ; 


XYII. 


VIL 


On Some Points in Dynamical 
Geology; 




WW 


On Limestone, Dolomites and 
Gypsums; 


XYIII. 


IX. 


The Chemistry of Natural Waters; 


XIX. 


X. 


Petroleum, Asphalt, Pyroschists 






and Coal; 


XX. 


XI. 


Granites and Granitic Yein- 
stones; 





The Origin of Metalliferous De- 
posits ; 

The Geognosy of the Appala- 
chians and the Origin of Crys- 
talline Rocks; 

The Geology of the Alps; 

History of the Names Cambrian 
and Silurian in Geology ; 

Theory of Chemical Changes and 
Equivalent Yolumes; 

The Constitution and Equiva- 
lent Yolume of Mineral 
Species; 

Thoughts on Solution and the 
Chemical Process ; 

On the Objects and Method of 
Mineralogy ; 

The Theory of Types in Chem- 
istry. 

Appendix and Index. 



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XX ADVERTISEMENTS, 



A NEW BASIS FOR CHEMISTRY. 



A CHEMICAL PHILOSOPHY 



BY 



THOMAS STERRY HUNT, m.a., ll. d., 



Author of " Chemical and Geological Essays," •• Mineral Physiology 
and Physiography," " Systematic Mineralogy," etc. 



FOURTH EDITION. 



PRICE, &2.00. 



TABLE OF 1 CONTENTS. 



1. Introduction. \ 

II. Nature of the Chemical Process. 
III. Genesis of the Chemical Ele 
ments. 
IV Gases, Liquids and Solids, 
V. The Law of Numbers. 
VI. Equivalent Weights. 
VII. Hardness ana Chemi 
cal Indifference. 



VIII. The Atomic Hypothesis. 
IX. The Law of Volumes. 
X. Metamorphosis in Chemistry. 
XI. The Law of Densities. 
XII. Historical Retrospect. 
XIII. Conclusions. 
XIV. Supplement. 
Appendix and Index. 



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ADVERTISEMENTS. XXI 



MINERAL 

PHYSIOLOGY AND PHYSIOGRAPHY. 

A SECOND SERIES OF 

CHEMICAL AND GEOLOGICAL ESSAYS, 

WITH 

A GENERAL INTRODUCTION. 

BT 

THOMAS STERRY HUNT, m.a.,ll.d. 

Author of "Chemical and Geological Essays," "A New Basis for 
Chemistry," ".Systematic Mineralogy," etc 



SECOND EDITION. REVISED AND ENLARGED. 

PRICE, $5.00. 



TABLE OF COITTEaSTTS- 
Pbefacb. 

Chapter I.— Nature in Thought and Language. 
Chapter II.— The Order of the Natural Sciences. 

Chapter III.— Chemical and Geological Relations of the Atmosphere. 
Chapter IV.— Celestial Chemistry from the Time of Newton. 
Chapter V.— The Origin of Crystalline Rocks. 

Chapter VI.— The Genetic History of Crystalline Rocks. 
Chapter VII.— The Decay of Crystalline Rocks. 
Chapter VIII.— A Natural System in Mineralogy, with a Classification of Silicates. 
Chapter IX.— History of Pre-Cambrian Rocks. 

Chapter X.— The Geological History of Serpentine, with Studies of Pre- 
Cambrian Rocks. 
Chapter XI.— The Taconic Question in Geology. 
Appendix and Index. 



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XXII 



ADVERTISEMENTS. 



SYSTEMATIC MINERALOGY 



BASED ON A 



NATURAL CLASSIFICATION 

WITH A GENERAL INTRODUCTION. 



THOMAS STERRY HUNT, m.a., ll.d. ( 

Author of "Chemical and Geological Essays," "Mineral Physiology and 
Physiography," "A New Basis for Chemistry," etc. 
SECOND EDITION. 



BOUND IN CLOTH. PRICE $5.00. 



The aim of the author in the present treatise has been to reconcile the 
rival and hitherto opposed Chemical and Natural History methods in Min- 
eralogy, and to constitute a new system of classification, which is "at the 
same time Chemical and Natural Historical,' 'or, in the words of the preface, 
" to observe a strict conformity to chemical principles, and at the same time 
to retain all that is valuable in the Natural History method ; the two 
opposing schools being reconciled by showing that when rightly under- 
stood, chemical and physical characters are really dependent on each other, 
and present two aspects of the same problem which can never be solved but 
by the consideration of both." He has, moreover, devised and adopted a 
Latin nomenclature and arranged the mineral kingdom in classes, orders, 
genera and species, the designations of the lat:er being binomial, 



TABLE OF CONTENTS. 



Chapter I. The Relations of Mineralogy; 
II. Mineralogical Systems; 

III. First Principles in Chemis- 

try; 

IV. Chemical Elements and No- 

tation; 
V. Specific Gravity; 
VI. The Coefficient of Mineral 

Condensation; 
TIL The Theory of Solution; 
Till. Relations of Condensation to 
Hardness and Insolubility; 
IX. Crystallization and its Rela- 
tions; 



Chapter X. The Constitution of Mineral 
Species; 
XL A Ne w Mineralogical Classi" 
ficatioa; 
XII. Mineralogical Nomencla- 
ture; 
XIII. Synopsis of Mineral Species; 
XIT. The Metallaceous Class; 
XT. The Halidaceous Class. 
XTI. The Oxydaceous Class; 
XTII. The Pyricaustaceous Class; 
XVIII. Mineral History of Waters; 
General Index; 
Index of Names of Minerals. 



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